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Extraction of lithium, rubidium and cesium from lithium porcelain stone
Hydrometallurgy 191 (2020) 105233
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Extraction of lithium, rubidium and cesium from lithium porcelain stone a,b
Jinliang Wang a b
a,b,⁎
, Huazhou Hu
a,b
, Kaiqi Wu
T
Faculty of Materials Metallurgy and Chemistry, Jiangxi University of Science and Technology, Ganzhou 341000, China Jiangxi Provincial Key Laboratory of Flash Green Development and Recycling, Jiangxi University of Science and Technology, Ganzhou 341000, China
ARTICLE INFO
ABSTRACT
Keywords: Lithium porcelain stone Alkali metals Lithium Rubidium Cesium
The anticipated increase in demand for Li, Rb, and Cs raises growing concern due to the expected shortage in alkali metals available for extraction from ores. In this study, two kinds of additives (Na2SO4 and CaCl2) were used to selectively extract alkali metals from lithium porcelain stone. HSC modelling was used to simulate the process of roasting the ore in the presence of these additives, and to predict the temperature effect in addition to the ore/Na2SO4/CaCl2 mass ratio. Under the optimized conditions, the Li, Na, K, Rb, and Cs extraction efficiencies were 98.70, 49.80, 37.90, 97.27, and 98.40%, respectively. The optimal roasting conditions were found to be an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2 with heating at 850 °C for 60 min, while the optimal water leaching conditions were a liquid/solid ratio of 1/1 and a stirring speed of 300 rpm at room temperature for 60 min. Upon analysis of the calcined and leaching residues, the extraction mechanism was demonstrated to involve a highly chemoselective ion-exchange process between the alkali metals present in the ore, the Na+ in Na2SO4, and the Ca2+ in CaCl2. After this process, the concentrations of the alkali metals in lixivium were approximately 3.43 g/L Li, 25.50 g/L Na, 11.00 g/L K, 1.35 g/L Rb, and 0.56 g/L Cs, thereby indicating the feasibility of this process.
1. Introduction With the introduction of the international new energy development plan, lithium has become one of the key supporting international energy sources. In this context, lithium sulfate and lithium carbonate are important basic raw materials for the development of lithium as a new energy source, and demand for the latter is projected to increase by ~60% (i.e., from 102,000 to 162,000 tons of lithium carbonate equivalent) in the next 5 years. Indeed, its application in lithium-ion batteries drives a large percentage (~40,000 tons) of this growth (Siame and Pascoe, 2011; Luong et al., 2013). Two worldwide economic sources of lithium exist, namely brines and hard rock ores. More specifically, lithium-containing brines make up 66% of the world's lithium resources, with pegmatites making up 26% and sedimentary rocks making up 8% of this total (Meshram et al., 2014). In 2011, 85% of the world lithium production from ores originated from Australia's Greenbushes spodumene pegmatite deposit as a lithium mineral concentrate, and this mineral is exported to China to produce lithium carbonate and lithium hydroxide (Peiro et al., 2011). To meet the growing demand for Li, lithium-bearing minerals, including spodumene (Chen et al., 2011), lepidolite (Vieceli et al., 2017; Meshram et al., 2014), and zinnwaldite (Jandová et al., 2010), have also drawn attention in recent years. However,
⁎
few studies have focused on the extraction of lithium from lithium porcelain stone because of its low Li content (i.e., an average Li2O content of 1.00%) (Zhou et al., 2015; He et al., 2014; Platova et al., 2014). Indeed, the extraction of lithium from lithium porcelain stone is challenging, and so it is not considered an efficient utilization of resources. The exploration of lithium porcelain stone in China is mainly concentrated in the Jiangxi province, with the stones being mainly distributed in the Yifeng, Fengxin, and Shanggao regions. The mineralization process of lithium and rare metals is rather complicated, with magmatism and tectonic activities playing leading roles. As of 2010, estimated resource reserves of 81.6 million tons exist (Zhong et al., 2018), and a total of 24.37 million tons of lithium porcelain stone have been identified in the Dang Tian mining area in Yifeng County, along with 0.4 million tons of Li2O (Wu et al., 2016). At present, a small number of lithium porcelain stones have been found to yield lithium through the priority extraction of lepidolite from the ore. The majority of this ore is cheap, and goes directly to glass or ceramics factories as a raw material (Hermawan et al., 2009; Liu and Liu, 2014; Maslennikova et al., 1993). In one respect, the process of refining lepidolite from lithium porcelain stone produces a large number of tailings, although excess energy and manpower are consumed, thereby generating high production costs. Likewise, the recovery rate of lithium during lepidolite extraction from lithium porcelain stone reaches only 70–80% (Liu and Liu, 2014), while the
Corresponding author at: Faculty of Materials Metallurgy and Chemistry, Jiangxi University of Science and Technology, Ganzhou 341000, China. E-mail address: [email protected] (H. Hu).
https://doi.org/10.1016/j.hydromet.2019.105233 Received 26 June 2019; Received in revised form 29 November 2019; Accepted 17 December 2019 Available online 19 December 2019 0304-386X/ © 2019 Elsevier B.V. All rights reserved.
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Table 1 Result of elemental analysis of the lithium porcelain stone.
12000 1-SiO2 2-NaAlSi3O8 3-K ( Li, Al )3 ( Si, Al )4O10 ( F, OH )2 4-K1.76Na0.12Rb0.10Ca0.02Li3.14Al4.14Si6.72O20.94F3.06
10000
Element
Li
Na
K
Rb
Cs
Ca
Fe
SiO2
Al2O3
F
wt. (%)
0.48
1.90
4.45
0.19
0.08
0.35
0.05
70.31
18.00
1.02
Counts
8000
porcelain stone was crushed in a jaw crusher, then ground in a sampling machine and screened to a particle size of < 150 μm. The X-ray powder diffraction (XRD) pattern shown in Fig. 1 indicates that the raw material was primarily composed of SiO2, NaAlSi3O8, K(Li, Al)3(Si, Al)4O10 (F, OH)2, and K1.76Na0.12Rb0.10Ca0.02Li3.14Al4.14Si6.72O20.94F3.06. In addition, Fig. 2 shows a smooth surface and a uniform particle size. The results of chemical analysis of the raw material are listed in Table 1. The Na2SO4, CaCl2, and other reagents employed were of analytical grade and were purchased from the Xilong Chemical Co., Ltd. (Guangdong, China). Atomic absorption spectrometry (AAS, TAS990, PuXi China) was used to determine the contents of Li, Na, K, Rb, and Cs using standard procedures. Inductively coupled plasma optical emission spectroscopy (ICP-OES, IRIS Intrepid II XSP, Thermo Electron Corporation China) was employed to determine the contents of other elements present in the lixivium, and the pH was determined using a pH meter (ST3100, OHAUS America). The various samples, the calcine, and the residue were analyzed by XRD (Empyrean, PANalytical Holland) and scanning electron microscopy-energy dispersive spectroscopy (SEM-EDS, MLA650F, Bruker Germany).
6000
2
4000
2000
2
4 3
4 3 2
4 4 3 3
2 2
0 20
40
60
80
Position [°2Theta] Fig. 1. XRD pattern of the raw material lithium porcelain stone.
recovery rate of lithium during lepidolite extraction is 80–95% (Yan et al., 2012b; Luong et al., 2013), and that of comprehensive lithium is only 60–75%. As a result, a large portion of lithium resources are wasted, and as a whole, this extraction process is complicated and has serious drawbacks. It has been reported that salt roasting with Na2SO4 and CaCl2 followed by water leaching can efficiently extract Li, Rb, and Cs from lepidolite (Yan et al., 2012a). As such, we wished to investigate the use of two different additives (i.e., Na2SO4 and CaCl2) for enhanced extraction. However, the contents and phases of low grade alkali metals are known to differ significantly between lepidolite and lithium porcelain stone. We therefore aim to optimize the conditions for the recovery of alkali metals directly from lithium porcelain stone (Li2O ~1.0%, China) using Na2SO4 or CaCl2 roasting followed by water leaching to determine which factors affect its extraction. Specific attention is given to the highly chemoselective ion-exchange process between alkali metals present in the ore, Na+ from Na2SO4, and Ca2+ in CaCl2.
2.2. Experimental apparatus and procedure The procedure employed for the extraction of alkali metals from the lithium porcelain stone is outlined in Fig. 3. More specifically, a sealed tube furnace (SK-1600 °C, Zhonghuan China) was used for the roasting tests, and an asynchronous digital constant speed electric mixer (RHYG4S, Wanghua China) was used for the leaching experiments. In the roasting and leaching processes, a sample of the ore (20 g) was initially added to a 100 mL corundum crucible and mixed with Na2SO4 and CaCl2 for roasting. The resulting calcine was then added under stirring at 300 rpm to a 250 mL beaker with a liquid/solid (L/S) ratio of 2/1, and subjected to leaching for 60 min at room temperature. During the leaching process, the alkali metal concentrations in the lixivium were analyzed periodically to determine the optimal leaching time. The solution was then filtered to separate the lixivium from the residue, and the latter was subjected to leaching once again by washing to reduce loss of the residue. The leaching rates of the alkali metals were calculated based on changes in the concentration of the ore in the lixivium using the following formula:
2. Experimental 2.1. Materials and analysis The lithium porcelain stone employed herein was provided by the Feiyu New Energy Technology Co., Ltd. (Jiangxi, China). Initially, the lithium
Fig. 2. SEM pattern of the raw material lithium porcelain stone. 2
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Ore, Na2SO4 and CaCl2
Roasting
existence of SiO2 and Al2O3 in the ore instigates the forward reaction (1), which is shown in reaction (5). Reactions (5–8, 13–16) can be orchestrated so that Na2SO4/CaCl2 can interact with alkali metal oxides in the lithium porcelain stone in the presence of SiO2 and Al2O3. With the increase of atomic number, the reaction trends of K, Rb, and Cs also increases. Compared to Li2O, the gibbs free energy of the oxides of K, Rb, and Cs reacting with Na2SO4/CaCl2 is significantly smaller, so it can be deduced that the oxides of K, Rb, and Cs are easier to sulfate/ chlorinate than Li2O. Compared with Na2SO4, the gibbs free energy of Na2SO4/CaCl2 reacting with the ore is significantly smaller, so it was easier to chlorinate than sulfate the ore under the corresponding temperature. However, it needs to be further explored which reaction happens during the roasting process of lithium porcelain stone, and which reaction plays the leading role.
Water
Leaching solution
Leaching Residue Water
Washing Residue
3.2. Effect of ore/Na2SO4 mass ratio on alkali metals leaching rates
Fig. 3. Schematic of procedure to treat lithium porcelain stone.
=
The effect of the ore/Na2SO4 mass ratio was then examined between 1/0 and 1/0.4 g/g, while other parameters remained fixed, i.e., an ore/ CaCl2 mass ratio of 1/0.1 g/g, a temperature of 950 °C, and a time of 60 min. Each calcine was leached for 60 min at room temperature, with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. The relationship between the alkali metal leaching rate and the ore/Na2SO4 mass ratio is therefore outlined in Fig. 5, while the XRD patterns of the calcines obtained at different ore/Na2SO4 mass ratios are given in Fig. 6. As indicated, the alkali metal leaching rates gradually increased upon increasing the ore/Na2SO4 mass ratio. More specifically, as the amount of Na2SO4 increased, ion-exchange between the alkali metal and the Na+ present in Na2SO4 was promoted, and at an ore/Na2SO4 mass ratio of 1/0.2, the Li leaching rate reached a maximum of 97.63% and maintained a steady value. The K, Rb, and Cs leaching rates were also found to increase upon increasing the ore/Na2SO4 mass ratio. In addition, as shown in Fig. 4, the leaching of K, Rb, and Cs is more facile than in the case of Li, and a small amount of alkali metal salts may be volatized at 950 °C. The optimum ore/ Na2SO4 mass ratio for roasting was therefore considered to be 1/0.2. Furthermore, the data presented in Fig. 6 indicate that in the absence of Na2SO4, alkali metal leaching did not occur, and indeed, the XRD pattern of this calcine showed no clear change compared to the raw material. Upon increasing the quantity of Na2SO4, the phase of the calcine clearly transformed from SiO2 and NaAlSi3O8 to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4. The phases of Li, Rb, and Cs in the calcine were not characterized due to their low contents.
c V + c1 V 1 × 100 m 0 w0 + m1 w1
where φ is the leaching rate, %; m0 indicates the mass of the initial material, g; w1 indicates the relevant element mass ratio of the added additives, g; m1 indicates the mass of the added additives, g; w0 indicates the relevant element mass ratio of the initial material, g; c is the concentration of the relevant element in the filtrate after the leaching process, g/L; V is the volume of the corresponding filtrate after the leaching process, L; c1 is the concentration of the relevant element in the filtrate after the washing process, g/L; and V1 is the volume of the corresponding filtrate after the washing process, L. 3. Results and discussion 3.1. Thermodynamic modelling using HSC Due to the lack of thermodynamic data on lithium porcelain stone, only the reaction free energy changes between simple compounds can be used to investigate the reaction Gibbs free energy changes of Na2SO4 and CaCl2 to the main components of lithium porcelain stone at atmospheric pressure. Thus, the corresponding equations are listed in Table 2, and thermodynamic modelling was performed using the HSC program (Outotec, 2011) to predict which factors control the extraction of lithium during the roasting process. The relationship between temperature T and the reaction Gibbs free energy change △rGmθ is shown in Fig. 4. As shown in Fig. 4, the pattern of T shows that the △rGmθ of reaction (1) is greater than zero in the given temperature range, indicating that Na2SO4 cannot directly interact with Li2O. However, the
3.3. Effect of ore/CaCl2 mass ratio on alkali metals leaching rates The influence of the ore/CaCl2 mass ratio on the alkali metal leaching rates and the calcine phase was then examined as outlined in
Table 2 Reactions of the roasting process. No.
Reactions
△rGmθ~T kJ/mol
(1) (2) (3) (4) (5) (6) (7) (8) (9) (10) (11) (12) (13) (14) (15) (16)
Na2SO4 + Li2O = Li2SO4 + Na2O Na2SO4 + K2O=K2SO4 + Na2O Na2SO4 + Rb2O = Rb2SO4 + Na2O Na2SO4 + Cs2O=Cs2SO4 + Na2O Na2SO4 + Li2O + 6SiO2 + Al2O3 = Li2SO4 + 2NaAlSi3O8 Na2SO4 + K2O + 6SiO2 + Al2O3 = K2SO4 + 2NaAlSi3O8 Na2SO4 + Rb2O + 6SiO2 + Al2O3 = Rb2SO4 + 2NaAlSi3O8 Na2SO4 + Cs2O + 6SiO2 + Al2O3 = Cs2SO4 + 2NaAlSi3O8 CaCl2 + Li2O=CaO + 2LiCl CaCl2 + K2O=CaO + 2KCl CaCl2 + Rb2O=CaO + 2RbCl CaCl2 + Cs2O=CaO + 2CsCl CaCl2 + Li2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2LiCl CaCl2 + K2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2KCl CaCl2 + Rb2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2RbCl CaCl2 + Cs2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2CsCl
131.31–0.0029 T −106.171–0.005 T −127.982 + 0.002 T −126.325 + 0.00956 T −101.279–0.0002 T −338.76–0.0024 T −360.57 + 0.0047 T −358.913 + 0.0122 T −58.183–0.0107 T −351.198–0.0008 T −371.97–0.0048 T −377.75–0.0148 T −147.346–0.0131 T −440.361–0.0015 T −461.361 + 0.0024 T −466.913 + 0.0125 T
3
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Fig. 4. Relationship between the temperature T and the reaction gibbs free energy △rGmθ (HSC 7.1, Outotec).
Li Na K Rb Cs
100
Leaching rate /%
80
60
40
20
Fig. 5. Effect of ore/Na2SO4 mass ratio on alkali metals leaching rates.
0 1:0.2:0
1:0.2:0.1
1:0.2:0.2
1:0.2:0.3
1:0.2:0.4
ore/Na2SO4/CaCl2 mass ratio
Fig. 7. Effect of ore/CaCl2 mass ratio on alkali metals leaching rates. 20000
3-CaSO4
1-SiO2 2-NaCl
4-(Na,Ca) Al (Si,Al)3O8 5-NaAlSi3O8
5 4
1
16000
44 155 3
2
1 44 155 3 1
12000
8000 15 3
2 5 5
5 4 2
ore/CaCl2=1/0.3 1
ore/CaCl2=1/0.2
2
5 4 2
5 5
1
2 5 5
1
5
1
ore/CaCl2=1/0.1
1
4000
5 155
Fig. 6. XRD pattern of the calcine with the change of ore/Na2SO4 mass ratio.
ore/CaCl2=1/0
0 10
Figs. 7 and 8. All other roasting parameters were maintained constant, i.e., an ore/Na2SO4 ratio of 1/0.2, a roasting temperature of 950 °C, and a time of 60 min. Each calcine was leached for 60 min at room temperature using an L/S ratio of 2/1 and a stirrer speed of 300 rpm. As outlined in Fig. 7, the ore/CaCl2 mass ratio was also found to a have a significant effect on the alkali metal leaching rates. More specifically, the Li, Rb, and Cs leaching rates increased from 35.00, 2.10, and 4.90%, respectively, at an ore/CaCl2 mass ratio of 1/0 to 98.65, 70.83, and 84.00%, respectively, at an ore/CaCl2 mass ratio of 1/0.2. Upon increasing the ore/CaCl2 mass ratio beyond 1/0.2, the alkali
20
30
40
50
60
70
80
90
Position [°2Theta] Fig. 8. XRD pattern of the calcine with the change of ore/CaCl2 mass ratio.
metal leaching rates reached a maximum and remained stable, thereby indicating an optimal ore/CaCl2 mass ratio of 1/0.2 for this process. In addition, as shown in Fig. 8, the main calcine phases obtained at different ore/CaCl2 mass ratios were fundamentally different, and the main phases of SiO2 and NaAlSi3O8 were converted to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4 upon increasing the ore/CaCl2 4
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Fig. 12. XRD pattern of calcine with change in roasting time.
Fig. 9. Effect of roasting temperature on alkali metals leaching rates.
Fig. 13. Effect of leaching time on metals leaching rates.
Fig. 10. XRD pattern of the calcine with the change of roasting temperature.
Fig. 14. Effect of L/S ratio on metals leaching rates.
Fig. 11. Effect of stirring speed on alkali metals leaching rates.
5
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Table 3 Result of elemental analysis of the lixivium g/L. Li
Na
K
Rb
Cs
Ca
Mg
Al
Fe
SO42−
Cl−
pH
3.430
25.500
11.000
1.350
0.560
0.248
0.030
0.015
0.012
35.540
41.825
6.500
Fig. 17. Water-cum-material balance flowsheet. Fig. 15. Effect of temperature on metals leaching rates
Table 4 Result of elemental analysis of the residue. Element
Li
Na
K
Rb
Cs
Ca
Fe
SiO2
Al2O3
F
wt. (%)
0.02
0.79
1.48
0.03
/
6.05
0.03
58.37
14.94
0.79
5000
1
4000
Counts
1-SiO2 2-NaCl 3-CaSO4 4-(Na,Ca) Al (Si,Al)3O8 5-NaAlSi3O8
5 4
3000
2000
1000
5
2 5 45 3 5 14 4
Fig. 16. Effect of stirring speed on metals leaching rates.
2 1
3
1
2 1
1
0
mass ratio. In particular, the characteristic peak strength of (Na,Ca)Al (Si,Al)3O8 increased, while that of SiO2 decreased.
20
40
60
80
Position [°2Theta]
3.4. Effect of roasting temperature on alkali metals leaching rates
Fig. 18. XRD pattern of the calcine.
roasting temperature was considered to be 850 °C. In addition, Fig. 10 shows that the main calcine phases differed upon variation in the roasting temperature, and the main phases of SiO2 and NaAlSi3O8 were converted to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4, while the characteristic peak strength of SiO2 decreased with increasing temperature.
The effect of the roasting temperature was then examined, while the other parameters were fixed at an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/ 0.2 and a roasting time of 60 min. Each calcine was leached for 60 min at room temperature with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. The relationship between the alkali metal leaching rate and the roasting temperature was then examined (see Fig. 9), where it became apparent that the roasting temperature had a significant effect on the leaching rates. More specifically, upon increasing the roasting temperature from 750 to 850 °C, the Li, Rb, and Cs leaching rates reached their maximum values of 98.70, 97.27, and 98.40%, respectively. However, upon increasing the temperature further, the Rb, Cs, and K leaching rates decreased once again. It is expected that since chlorine has a lower boiling point, the roasting temperature of chlorine is lower than that of sulfate (Yan et al., 2012a), and so the optimal
3.5. Effect of roasting time on alkali metals leaching rates The effect of the roasting time was then examined with the other parameters fixed at an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2 and 850 °C. Each calcine was then leached for 60 min at room temperature with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. Thus, as shown in Fig. 11, the alkali metal leaching rates increased upon increasing the 6
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8000
6000
Counts
namely SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4. Furthermore, the characteristic peak strengths of NaAlSi3O8 and (Na,Ca)Al (Si,Al)3O8 increased upon lengthening the roasting time, while that of SiO2 decreased.
1-SiO2 2-CaSO4 3-(Na,Ca) Al (Si,Al)3O8 4-NaAlSi3O8
21
3.6. Effect of leaching factors on alkali metals leaching rates
4 3 4000
4 13
2000
4
2
22 11314
4 1
1
The effects of the leaching time and the L/S ratio on the leaching recovery of the metals were then investigated under initial conditions of 100 g calcine, a leaching temperature of room temperature, and a stirrer speed of 300 rpm. As presented in Fig. 13, the obtained results indicated that a leaching time > 60 min had only a slight influence on the extraction efficiency of the alkali metals, while an L/S ratio of 1/1 (see Fig. 14) tended to give an extraction efficiency 5–15% higher than that obtained with a ratio of 0.8/1. More specifically, the Li, Na, K, Rb, and Cs extraction efficiencies peaked at 98.10, 49.80, 37.90, 97.15, and 98.24% with an L/S ratio of 1/1 and a leaching time of 60 min. Although a low L/S ratio will increase the lixivium concentration, it will render agitation difficult and lead to insufficient leaching (Luong et al., 2014; Hien-Dinh et al., 2015), and so an L/S ratio of 1/1 was considered to be optimal. Interestingly, the leaching temperature was found to have no significant effect on the alkali metal leaching rates, and due to the low L/S ratio, a stirring speed of at least 300 rpm was required to facilitate dissolution. Therefore, using the optimized leaching
1
0 20
40
60
80
Position [°2Theta] Fig. 19. XRD pattern of the leaching residue.
roasting time up to 60 min. However, beyond this time, the Rb, Cs, and K leaching rates decreased slightly, and so a roasting time of 60 min was considered optimal. In addition, Fig. 12 shows the XRD patterns of the calcines obtained at different roasting times, where it is apparent that the main calcine phases were fundamentally similar throughout,
Fig. 20. SEM-EDS pattern of the calcine. 7
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Fig. 21. SEM-EDS pattern of the leaching residue.
conditions, the resulting lixivium was found to contain ~3.43 g/L Li, ~25.50 g/L Na, ~11.00 g/L K, ~1.35 g/L Rb, and ~0.56 g/L Cs, as presented in Table 3. (See Figs. 15 and 16.) According to the optimal process conditions obtained in the previous experiments, the water-cum-material balance flowsheet was carried out on the process of leaching alkali metlas from lithium porcelain stone. The result is shown in Fig. 17.
NaCl. Meanwhile, calcium chloride facilitates this process. From Figs. 20 and 21, the soluble salt dissolves and the surface of leaching residue is sparse, porous, and dispersed after the calcine leached. This also indicates that the soluble salts of alkali metals were still leached, and then in the roasting process fully converted. After leaching, the contents of Na, K, S, Cl and the particle size of calcine clearly decreased, but the contents of O, Si, and Al did not leach during the water leaching process.
3.7. Analysis of calcine and residue
4. Conclusions
The roasting and leaching experiments were conducted under the optimum technological conditions of an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2, roasting temperature of 850 °C, roasting time of 60 min, solid/liquid ratio of 1/1, leaching at room temperature for 60 min, and stirring speed of 300 rpm. The calcine and leaching residue under the optimum technological conditions were analyzed by ICP-MS, XRD and SEM-EDS, respectively. The results are shown in Table 4 and Figs. 18~21. As shown in Table 4, the components of Li, Rb, Cs in the ore had been nearly complete leached and most of the F remained in residue. In addition, the leaching residue/lithium porcelain stone mass ratio was approximately 1.2. As shown in Fig. 18, new phases of NaCl, CaSO4, (Na,Ca)Al(Si,Al)3O8 were formed in the roasting process. Compared with Fig. 19, the most soluble salts were followed by water leaching. The extraction mechanism was found to be high-chemoselective ion-exchange between alkali metals in ore, and Na+ in Na2SO4/
We herein investigated the effects of the Na2SO4 and CaCl2 roasting additives on the selective water leaching of alkali metals from lithium porcelain stone. This procedure is expected to be a relatively cost-effective and straightforward method for the extraction of lithium, and an alternative to current industrial processes. Under the optimized roasting conditions of an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2, a temperature of 850 °C, and a time of 60 min, in addition to optimized leaching conditions of a liquid/solid ratio of 1:1, a temperature of 25 °C, a stirrer speed of 300 rpm, and a time of 60 min, the maximum leached contents of Li, Na, K, Rb, and Cs were 98.10, 49.80, 37.90, 97.15, and 98.24% respectively. Likewise, the concentrations of the alkali metals in the resulting lixivium were ~3.43 g/L Li, ~25.50 g/L Na, ~11.00 g/ L K, ~1.35 g/L Rb, and ~0.56 g/L Cs, thereby indicating the feasibility of this process. Furthermore, based on HSC modelling and X-ray 8
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diffraction measurements, the extraction mechanism was found to involve a highly chemoselective ion-exchange process between the alkali metals present in the ore, the Na+ present in Na2SO4, and the Ca2+ present in CaCl2. These results are of importance due to the anticipated increase in demand for Li, Rb, and Cs accompanied by an expected shortage in alkali metals available for extraction from ores.
China Non-Metallic Miner. Ind. 6, 41–43 (In Chinese). Hermawan, A., Ohuchi, T., Fujimoto, N., et al., 2009. Manufacture of composite board using wood prunings and waste porcelain stone[J]. Journal of Wood Science 55 (1), 74–79. Hien-Dinh, T.T., Luong, V.T., Gieré, R., Tran, T., 2015. Extraction of lithium from lepidolite via iron sulphide roasting and water leaching. Hydrometallurgy 153, 154–159. Jandová, J., Dvořák, P., Hong, N.V., 2010. Processing of zinnwaldite waste to obtain Li2CO3. Hydrometallurgy 103 (1), 12–18. Liu, Y.L., Liu, J., 2014. The flotation process of lepidolite in Jiangxi province in China. Adv. Mater. Res. 4, 1033–1034. Luong, V.T., Dong, J.K., An, J.W., Kim, M.J., Tran, T., 2013. Factors affecting the extraction of lithium from lepidolite. Hydrometallurgy 134 (3), 54–61. Luong, V.T., Dong, J.K., An, J.W., Dao, D.A., Kim, M.J., Tran, T., 2014. Iron sulphate roasting for extraction of lithium from lepidolite. Hydrometallurgy 141 (2), 8–16. Maslennikova, G.N., Platov, Y.T., Zhekisheva, S.Z., 1993. Porcelain stone - a nontraditional type of mineral raw material. Glas. Ceram. 50 (11−12), 472–475. Meshram, P., Pandey, B.D., Mankhand, T.R., 2014. Extraction of lithium from primary and secondary sources by pre-treatment, leaching and separation: a comprehensive review. Hydrometallurgy 150, 192–208. Peiro, L.T., Mendez, G.V., Ayres, R.U., 2011. Lithium: sources, production, uses, and recovery outlook. JOM 65 (8), 986–996. Platova, R.A., Platov, Y.T., 2014. Biochemical Method of Iron Removal and its Effect on the Color Characteristics of Porcelain Stone. Glass & Ceramics 71 (1-2), 28–34. Outotec, 2011. HSC Chemistry 7.1. Siame, E., Pascoe, R.D., 2011. Extraction of lithium from micaceous waste from China clay production. Miner. Eng. 24 (14), 1595–1602. Vieceli, N., Nogueira, C.A., Pereira, M.F.C., Dias, A.P.S., Durão, F.O., Guimarães, C., Margarido, F., 2017. Effects of mechanical activation on lithium extraction from a lepidolite ore concentrate. Miner. Eng. 102, 1–14. Wu, X.M., Zhou, M.J., Luo, X.C., Zhou, J.T., 2016. The metallogenic conditions and prospecting potential of lithium and rare metals in northwestern Jiangxi. East China Geol. 4 (37), 275–283. Yan, Q.X., Li, X.H., Wang, Z.X., Wu, X.F., Guo, H.J., Hu, Q.Y., Peng, W.J., Wang, J.X., 2012a. Extraction of valuable metals from lepidolite. Hydrometallurgy 117-118 (4), 116–118. Yan, Q.X., Li, X.H., Yin, Z.L., Wang, Z.X., Guo, H.J., Peng, W.J., Hu, Q.Y., 2012b. A novel process for extracting lithium from lepidolite. Hydrometallurgy 121-124, 54–59. Zhong, B., Li, H.Z., Liu, J.Y., Wang, J.Q., Nan, J.X., Wu, J.F., 2018. A method for extracting lithium directly from the raw material of lithium porcelain stone. CN Patent 109055723A. Zhou, J.E., Liu, K., Dong, W.X., Bao, Q.F., Zhao, T.G., Wang, Y.Q., 2015. Effects of CaOLi2O-K2O-Na2O fluxing agents on the properties of porcelain ceramic tiles. Key Eng. Mater. 655, 258–262.
Declaration of Competing Interest We declare that we have no financial and personal relationships with other people or organizations that can inappropriately influence our work, there is no professional or other personal interest of any nature or kind in any product, service and/or company that could be construed as influencing the position presented in, or the review of, the manuscript entitled “Extraction of lithium, rubidium and cesium from lithium porcelain stone”. Acknowledgements This work was financially supported by the National Natural Science Foundation of China (No. 51564018 and No. 51764018), the Innovation Team Support Program of Jiangxi University of Science and Technology (3200826514), the Jiangxi Provincial Key Laboratory of Flash Green Developmet and Recycling (20193BCD40019), and the Program of Qingjiang Excellent Young Talents, Jiangxi University of Science and Technology. References Chen, Y., Tian, Q.Q., Chen, B.Z., Shi, X.C., Liao, T., 2011. Preparation of lithium carbonate from spodumene by a sodium carbonate autoclave process. Hydrometallurgy 109 (1), 43–46. He, M.S., Huang, X., Zou, G.S., 2014. Characteristics, Utilization Status and Suggestions of Lithium Containing Porcelain Stone (Soil) Mineral Resources in Jiangxi Province.
9
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Extraction of lithium, rubidium and cesium from lithium porcelain stone a,b
Jinliang Wang a b
a,b,⁎
, Huazhou Hu
a,b
, Kaiqi Wu
T
Faculty of Materials Metallurgy and Chemistry, Jiangxi University of Science and Technology, Ganzhou 341000, China Jiangxi Provincial Key Laboratory of Flash Green Development and Recycling, Jiangxi University of Science and Technology, Ganzhou 341000, China
ARTICLE INFO
ABSTRACT
Keywords: Lithium porcelain stone Alkali metals Lithium Rubidium Cesium
The anticipated increase in demand for Li, Rb, and Cs raises growing concern due to the expected shortage in alkali metals available for extraction from ores. In this study, two kinds of additives (Na2SO4 and CaCl2) were used to selectively extract alkali metals from lithium porcelain stone. HSC modelling was used to simulate the process of roasting the ore in the presence of these additives, and to predict the temperature effect in addition to the ore/Na2SO4/CaCl2 mass ratio. Under the optimized conditions, the Li, Na, K, Rb, and Cs extraction efficiencies were 98.70, 49.80, 37.90, 97.27, and 98.40%, respectively. The optimal roasting conditions were found to be an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2 with heating at 850 °C for 60 min, while the optimal water leaching conditions were a liquid/solid ratio of 1/1 and a stirring speed of 300 rpm at room temperature for 60 min. Upon analysis of the calcined and leaching residues, the extraction mechanism was demonstrated to involve a highly chemoselective ion-exchange process between the alkali metals present in the ore, the Na+ in Na2SO4, and the Ca2+ in CaCl2. After this process, the concentrations of the alkali metals in lixivium were approximately 3.43 g/L Li, 25.50 g/L Na, 11.00 g/L K, 1.35 g/L Rb, and 0.56 g/L Cs, thereby indicating the feasibility of this process.
1. Introduction With the introduction of the international new energy development plan, lithium has become one of the key supporting international energy sources. In this context, lithium sulfate and lithium carbonate are important basic raw materials for the development of lithium as a new energy source, and demand for the latter is projected to increase by ~60% (i.e., from 102,000 to 162,000 tons of lithium carbonate equivalent) in the next 5 years. Indeed, its application in lithium-ion batteries drives a large percentage (~40,000 tons) of this growth (Siame and Pascoe, 2011; Luong et al., 2013). Two worldwide economic sources of lithium exist, namely brines and hard rock ores. More specifically, lithium-containing brines make up 66% of the world's lithium resources, with pegmatites making up 26% and sedimentary rocks making up 8% of this total (Meshram et al., 2014). In 2011, 85% of the world lithium production from ores originated from Australia's Greenbushes spodumene pegmatite deposit as a lithium mineral concentrate, and this mineral is exported to China to produce lithium carbonate and lithium hydroxide (Peiro et al., 2011). To meet the growing demand for Li, lithium-bearing minerals, including spodumene (Chen et al., 2011), lepidolite (Vieceli et al., 2017; Meshram et al., 2014), and zinnwaldite (Jandová et al., 2010), have also drawn attention in recent years. However,
⁎
few studies have focused on the extraction of lithium from lithium porcelain stone because of its low Li content (i.e., an average Li2O content of 1.00%) (Zhou et al., 2015; He et al., 2014; Platova et al., 2014). Indeed, the extraction of lithium from lithium porcelain stone is challenging, and so it is not considered an efficient utilization of resources. The exploration of lithium porcelain stone in China is mainly concentrated in the Jiangxi province, with the stones being mainly distributed in the Yifeng, Fengxin, and Shanggao regions. The mineralization process of lithium and rare metals is rather complicated, with magmatism and tectonic activities playing leading roles. As of 2010, estimated resource reserves of 81.6 million tons exist (Zhong et al., 2018), and a total of 24.37 million tons of lithium porcelain stone have been identified in the Dang Tian mining area in Yifeng County, along with 0.4 million tons of Li2O (Wu et al., 2016). At present, a small number of lithium porcelain stones have been found to yield lithium through the priority extraction of lepidolite from the ore. The majority of this ore is cheap, and goes directly to glass or ceramics factories as a raw material (Hermawan et al., 2009; Liu and Liu, 2014; Maslennikova et al., 1993). In one respect, the process of refining lepidolite from lithium porcelain stone produces a large number of tailings, although excess energy and manpower are consumed, thereby generating high production costs. Likewise, the recovery rate of lithium during lepidolite extraction from lithium porcelain stone reaches only 70–80% (Liu and Liu, 2014), while the
Corresponding author at: Faculty of Materials Metallurgy and Chemistry, Jiangxi University of Science and Technology, Ganzhou 341000, China. E-mail address: [email protected] (H. Hu).
https://doi.org/10.1016/j.hydromet.2019.105233 Received 26 June 2019; Received in revised form 29 November 2019; Accepted 17 December 2019 Available online 19 December 2019 0304-386X/ © 2019 Elsevier B.V. All rights reserved.
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Table 1 Result of elemental analysis of the lithium porcelain stone.
12000 1-SiO2 2-NaAlSi3O8 3-K ( Li, Al )3 ( Si, Al )4O10 ( F, OH )2 4-K1.76Na0.12Rb0.10Ca0.02Li3.14Al4.14Si6.72O20.94F3.06
10000
Element
Li
Na
K
Rb
Cs
Ca
Fe
SiO2
Al2O3
F
wt. (%)
0.48
1.90
4.45
0.19
0.08
0.35
0.05
70.31
18.00
1.02
Counts
8000
porcelain stone was crushed in a jaw crusher, then ground in a sampling machine and screened to a particle size of < 150 μm. The X-ray powder diffraction (XRD) pattern shown in Fig. 1 indicates that the raw material was primarily composed of SiO2, NaAlSi3O8, K(Li, Al)3(Si, Al)4O10 (F, OH)2, and K1.76Na0.12Rb0.10Ca0.02Li3.14Al4.14Si6.72O20.94F3.06. In addition, Fig. 2 shows a smooth surface and a uniform particle size. The results of chemical analysis of the raw material are listed in Table 1. The Na2SO4, CaCl2, and other reagents employed were of analytical grade and were purchased from the Xilong Chemical Co., Ltd. (Guangdong, China). Atomic absorption spectrometry (AAS, TAS990, PuXi China) was used to determine the contents of Li, Na, K, Rb, and Cs using standard procedures. Inductively coupled plasma optical emission spectroscopy (ICP-OES, IRIS Intrepid II XSP, Thermo Electron Corporation China) was employed to determine the contents of other elements present in the lixivium, and the pH was determined using a pH meter (ST3100, OHAUS America). The various samples, the calcine, and the residue were analyzed by XRD (Empyrean, PANalytical Holland) and scanning electron microscopy-energy dispersive spectroscopy (SEM-EDS, MLA650F, Bruker Germany).
6000
2
4000
2000
2
4 3
4 3 2
4 4 3 3
2 2
0 20
40
60
80
Position [°2Theta] Fig. 1. XRD pattern of the raw material lithium porcelain stone.
recovery rate of lithium during lepidolite extraction is 80–95% (Yan et al., 2012b; Luong et al., 2013), and that of comprehensive lithium is only 60–75%. As a result, a large portion of lithium resources are wasted, and as a whole, this extraction process is complicated and has serious drawbacks. It has been reported that salt roasting with Na2SO4 and CaCl2 followed by water leaching can efficiently extract Li, Rb, and Cs from lepidolite (Yan et al., 2012a). As such, we wished to investigate the use of two different additives (i.e., Na2SO4 and CaCl2) for enhanced extraction. However, the contents and phases of low grade alkali metals are known to differ significantly between lepidolite and lithium porcelain stone. We therefore aim to optimize the conditions for the recovery of alkali metals directly from lithium porcelain stone (Li2O ~1.0%, China) using Na2SO4 or CaCl2 roasting followed by water leaching to determine which factors affect its extraction. Specific attention is given to the highly chemoselective ion-exchange process between alkali metals present in the ore, Na+ from Na2SO4, and Ca2+ in CaCl2.
2.2. Experimental apparatus and procedure The procedure employed for the extraction of alkali metals from the lithium porcelain stone is outlined in Fig. 3. More specifically, a sealed tube furnace (SK-1600 °C, Zhonghuan China) was used for the roasting tests, and an asynchronous digital constant speed electric mixer (RHYG4S, Wanghua China) was used for the leaching experiments. In the roasting and leaching processes, a sample of the ore (20 g) was initially added to a 100 mL corundum crucible and mixed with Na2SO4 and CaCl2 for roasting. The resulting calcine was then added under stirring at 300 rpm to a 250 mL beaker with a liquid/solid (L/S) ratio of 2/1, and subjected to leaching for 60 min at room temperature. During the leaching process, the alkali metal concentrations in the lixivium were analyzed periodically to determine the optimal leaching time. The solution was then filtered to separate the lixivium from the residue, and the latter was subjected to leaching once again by washing to reduce loss of the residue. The leaching rates of the alkali metals were calculated based on changes in the concentration of the ore in the lixivium using the following formula:
2. Experimental 2.1. Materials and analysis The lithium porcelain stone employed herein was provided by the Feiyu New Energy Technology Co., Ltd. (Jiangxi, China). Initially, the lithium
Fig. 2. SEM pattern of the raw material lithium porcelain stone. 2
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Ore, Na2SO4 and CaCl2
Roasting
existence of SiO2 and Al2O3 in the ore instigates the forward reaction (1), which is shown in reaction (5). Reactions (5–8, 13–16) can be orchestrated so that Na2SO4/CaCl2 can interact with alkali metal oxides in the lithium porcelain stone in the presence of SiO2 and Al2O3. With the increase of atomic number, the reaction trends of K, Rb, and Cs also increases. Compared to Li2O, the gibbs free energy of the oxides of K, Rb, and Cs reacting with Na2SO4/CaCl2 is significantly smaller, so it can be deduced that the oxides of K, Rb, and Cs are easier to sulfate/ chlorinate than Li2O. Compared with Na2SO4, the gibbs free energy of Na2SO4/CaCl2 reacting with the ore is significantly smaller, so it was easier to chlorinate than sulfate the ore under the corresponding temperature. However, it needs to be further explored which reaction happens during the roasting process of lithium porcelain stone, and which reaction plays the leading role.
Water
Leaching solution
Leaching Residue Water
Washing Residue
3.2. Effect of ore/Na2SO4 mass ratio on alkali metals leaching rates
Fig. 3. Schematic of procedure to treat lithium porcelain stone.
=
The effect of the ore/Na2SO4 mass ratio was then examined between 1/0 and 1/0.4 g/g, while other parameters remained fixed, i.e., an ore/ CaCl2 mass ratio of 1/0.1 g/g, a temperature of 950 °C, and a time of 60 min. Each calcine was leached for 60 min at room temperature, with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. The relationship between the alkali metal leaching rate and the ore/Na2SO4 mass ratio is therefore outlined in Fig. 5, while the XRD patterns of the calcines obtained at different ore/Na2SO4 mass ratios are given in Fig. 6. As indicated, the alkali metal leaching rates gradually increased upon increasing the ore/Na2SO4 mass ratio. More specifically, as the amount of Na2SO4 increased, ion-exchange between the alkali metal and the Na+ present in Na2SO4 was promoted, and at an ore/Na2SO4 mass ratio of 1/0.2, the Li leaching rate reached a maximum of 97.63% and maintained a steady value. The K, Rb, and Cs leaching rates were also found to increase upon increasing the ore/Na2SO4 mass ratio. In addition, as shown in Fig. 4, the leaching of K, Rb, and Cs is more facile than in the case of Li, and a small amount of alkali metal salts may be volatized at 950 °C. The optimum ore/ Na2SO4 mass ratio for roasting was therefore considered to be 1/0.2. Furthermore, the data presented in Fig. 6 indicate that in the absence of Na2SO4, alkali metal leaching did not occur, and indeed, the XRD pattern of this calcine showed no clear change compared to the raw material. Upon increasing the quantity of Na2SO4, the phase of the calcine clearly transformed from SiO2 and NaAlSi3O8 to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4. The phases of Li, Rb, and Cs in the calcine were not characterized due to their low contents.
c V + c1 V 1 × 100 m 0 w0 + m1 w1
where φ is the leaching rate, %; m0 indicates the mass of the initial material, g; w1 indicates the relevant element mass ratio of the added additives, g; m1 indicates the mass of the added additives, g; w0 indicates the relevant element mass ratio of the initial material, g; c is the concentration of the relevant element in the filtrate after the leaching process, g/L; V is the volume of the corresponding filtrate after the leaching process, L; c1 is the concentration of the relevant element in the filtrate after the washing process, g/L; and V1 is the volume of the corresponding filtrate after the washing process, L. 3. Results and discussion 3.1. Thermodynamic modelling using HSC Due to the lack of thermodynamic data on lithium porcelain stone, only the reaction free energy changes between simple compounds can be used to investigate the reaction Gibbs free energy changes of Na2SO4 and CaCl2 to the main components of lithium porcelain stone at atmospheric pressure. Thus, the corresponding equations are listed in Table 2, and thermodynamic modelling was performed using the HSC program (Outotec, 2011) to predict which factors control the extraction of lithium during the roasting process. The relationship between temperature T and the reaction Gibbs free energy change △rGmθ is shown in Fig. 4. As shown in Fig. 4, the pattern of T shows that the △rGmθ of reaction (1) is greater than zero in the given temperature range, indicating that Na2SO4 cannot directly interact with Li2O. However, the
3.3. Effect of ore/CaCl2 mass ratio on alkali metals leaching rates The influence of the ore/CaCl2 mass ratio on the alkali metal leaching rates and the calcine phase was then examined as outlined in
Table 2 Reactions of the roasting process. No.
Reactions
△rGmθ~T kJ/mol
(1) (2) (3) (4) (5) (6) (7) (8) (9) (10) (11) (12) (13) (14) (15) (16)
Na2SO4 + Li2O = Li2SO4 + Na2O Na2SO4 + K2O=K2SO4 + Na2O Na2SO4 + Rb2O = Rb2SO4 + Na2O Na2SO4 + Cs2O=Cs2SO4 + Na2O Na2SO4 + Li2O + 6SiO2 + Al2O3 = Li2SO4 + 2NaAlSi3O8 Na2SO4 + K2O + 6SiO2 + Al2O3 = K2SO4 + 2NaAlSi3O8 Na2SO4 + Rb2O + 6SiO2 + Al2O3 = Rb2SO4 + 2NaAlSi3O8 Na2SO4 + Cs2O + 6SiO2 + Al2O3 = Cs2SO4 + 2NaAlSi3O8 CaCl2 + Li2O=CaO + 2LiCl CaCl2 + K2O=CaO + 2KCl CaCl2 + Rb2O=CaO + 2RbCl CaCl2 + Cs2O=CaO + 2CsCl CaCl2 + Li2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2LiCl CaCl2 + K2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2KCl CaCl2 + Rb2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2RbCl CaCl2 + Cs2O + 2SiO2 + Al2O3 = CaAl2Si2O8 + 2CsCl
131.31–0.0029 T −106.171–0.005 T −127.982 + 0.002 T −126.325 + 0.00956 T −101.279–0.0002 T −338.76–0.0024 T −360.57 + 0.0047 T −358.913 + 0.0122 T −58.183–0.0107 T −351.198–0.0008 T −371.97–0.0048 T −377.75–0.0148 T −147.346–0.0131 T −440.361–0.0015 T −461.361 + 0.0024 T −466.913 + 0.0125 T
3
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Fig. 4. Relationship between the temperature T and the reaction gibbs free energy △rGmθ (HSC 7.1, Outotec).
Li Na K Rb Cs
100
Leaching rate /%
80
60
40
20
Fig. 5. Effect of ore/Na2SO4 mass ratio on alkali metals leaching rates.
0 1:0.2:0
1:0.2:0.1
1:0.2:0.2
1:0.2:0.3
1:0.2:0.4
ore/Na2SO4/CaCl2 mass ratio
Fig. 7. Effect of ore/CaCl2 mass ratio on alkali metals leaching rates. 20000
3-CaSO4
1-SiO2 2-NaCl
4-(Na,Ca) Al (Si,Al)3O8 5-NaAlSi3O8
5 4
1
16000
44 155 3
2
1 44 155 3 1
12000
8000 15 3
2 5 5
5 4 2
ore/CaCl2=1/0.3 1
ore/CaCl2=1/0.2
2
5 4 2
5 5
1
2 5 5
1
5
1
ore/CaCl2=1/0.1
1
4000
5 155
Fig. 6. XRD pattern of the calcine with the change of ore/Na2SO4 mass ratio.
ore/CaCl2=1/0
0 10
Figs. 7 and 8. All other roasting parameters were maintained constant, i.e., an ore/Na2SO4 ratio of 1/0.2, a roasting temperature of 950 °C, and a time of 60 min. Each calcine was leached for 60 min at room temperature using an L/S ratio of 2/1 and a stirrer speed of 300 rpm. As outlined in Fig. 7, the ore/CaCl2 mass ratio was also found to a have a significant effect on the alkali metal leaching rates. More specifically, the Li, Rb, and Cs leaching rates increased from 35.00, 2.10, and 4.90%, respectively, at an ore/CaCl2 mass ratio of 1/0 to 98.65, 70.83, and 84.00%, respectively, at an ore/CaCl2 mass ratio of 1/0.2. Upon increasing the ore/CaCl2 mass ratio beyond 1/0.2, the alkali
20
30
40
50
60
70
80
90
Position [°2Theta] Fig. 8. XRD pattern of the calcine with the change of ore/CaCl2 mass ratio.
metal leaching rates reached a maximum and remained stable, thereby indicating an optimal ore/CaCl2 mass ratio of 1/0.2 for this process. In addition, as shown in Fig. 8, the main calcine phases obtained at different ore/CaCl2 mass ratios were fundamentally different, and the main phases of SiO2 and NaAlSi3O8 were converted to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4 upon increasing the ore/CaCl2 4
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Fig. 12. XRD pattern of calcine with change in roasting time.
Fig. 9. Effect of roasting temperature on alkali metals leaching rates.
Fig. 13. Effect of leaching time on metals leaching rates.
Fig. 10. XRD pattern of the calcine with the change of roasting temperature.
Fig. 14. Effect of L/S ratio on metals leaching rates.
Fig. 11. Effect of stirring speed on alkali metals leaching rates.
5
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Table 3 Result of elemental analysis of the lixivium g/L. Li
Na
K
Rb
Cs
Ca
Mg
Al
Fe
SO42−
Cl−
pH
3.430
25.500
11.000
1.350
0.560
0.248
0.030
0.015
0.012
35.540
41.825
6.500
Fig. 17. Water-cum-material balance flowsheet. Fig. 15. Effect of temperature on metals leaching rates
Table 4 Result of elemental analysis of the residue. Element
Li
Na
K
Rb
Cs
Ca
Fe
SiO2
Al2O3
F
wt. (%)
0.02
0.79
1.48
0.03
/
6.05
0.03
58.37
14.94
0.79
5000
1
4000
Counts
1-SiO2 2-NaCl 3-CaSO4 4-(Na,Ca) Al (Si,Al)3O8 5-NaAlSi3O8
5 4
3000
2000
1000
5
2 5 45 3 5 14 4
Fig. 16. Effect of stirring speed on metals leaching rates.
2 1
3
1
2 1
1
0
mass ratio. In particular, the characteristic peak strength of (Na,Ca)Al (Si,Al)3O8 increased, while that of SiO2 decreased.
20
40
60
80
Position [°2Theta]
3.4. Effect of roasting temperature on alkali metals leaching rates
Fig. 18. XRD pattern of the calcine.
roasting temperature was considered to be 850 °C. In addition, Fig. 10 shows that the main calcine phases differed upon variation in the roasting temperature, and the main phases of SiO2 and NaAlSi3O8 were converted to SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4, while the characteristic peak strength of SiO2 decreased with increasing temperature.
The effect of the roasting temperature was then examined, while the other parameters were fixed at an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/ 0.2 and a roasting time of 60 min. Each calcine was leached for 60 min at room temperature with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. The relationship between the alkali metal leaching rate and the roasting temperature was then examined (see Fig. 9), where it became apparent that the roasting temperature had a significant effect on the leaching rates. More specifically, upon increasing the roasting temperature from 750 to 850 °C, the Li, Rb, and Cs leaching rates reached their maximum values of 98.70, 97.27, and 98.40%, respectively. However, upon increasing the temperature further, the Rb, Cs, and K leaching rates decreased once again. It is expected that since chlorine has a lower boiling point, the roasting temperature of chlorine is lower than that of sulfate (Yan et al., 2012a), and so the optimal
3.5. Effect of roasting time on alkali metals leaching rates The effect of the roasting time was then examined with the other parameters fixed at an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2 and 850 °C. Each calcine was then leached for 60 min at room temperature with an L/S ratio of 2/1 and a stirrer speed of 300 rpm. Thus, as shown in Fig. 11, the alkali metal leaching rates increased upon increasing the 6
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8000
6000
Counts
namely SiO2, NaAlSi3O8, (Na,Ca)Al(Si,Al)3O8, NaCl, and CaSO4. Furthermore, the characteristic peak strengths of NaAlSi3O8 and (Na,Ca)Al (Si,Al)3O8 increased upon lengthening the roasting time, while that of SiO2 decreased.
1-SiO2 2-CaSO4 3-(Na,Ca) Al (Si,Al)3O8 4-NaAlSi3O8
21
3.6. Effect of leaching factors on alkali metals leaching rates
4 3 4000
4 13
2000
4
2
22 11314
4 1
1
The effects of the leaching time and the L/S ratio on the leaching recovery of the metals were then investigated under initial conditions of 100 g calcine, a leaching temperature of room temperature, and a stirrer speed of 300 rpm. As presented in Fig. 13, the obtained results indicated that a leaching time > 60 min had only a slight influence on the extraction efficiency of the alkali metals, while an L/S ratio of 1/1 (see Fig. 14) tended to give an extraction efficiency 5–15% higher than that obtained with a ratio of 0.8/1. More specifically, the Li, Na, K, Rb, and Cs extraction efficiencies peaked at 98.10, 49.80, 37.90, 97.15, and 98.24% with an L/S ratio of 1/1 and a leaching time of 60 min. Although a low L/S ratio will increase the lixivium concentration, it will render agitation difficult and lead to insufficient leaching (Luong et al., 2014; Hien-Dinh et al., 2015), and so an L/S ratio of 1/1 was considered to be optimal. Interestingly, the leaching temperature was found to have no significant effect on the alkali metal leaching rates, and due to the low L/S ratio, a stirring speed of at least 300 rpm was required to facilitate dissolution. Therefore, using the optimized leaching
1
0 20
40
60
80
Position [°2Theta] Fig. 19. XRD pattern of the leaching residue.
roasting time up to 60 min. However, beyond this time, the Rb, Cs, and K leaching rates decreased slightly, and so a roasting time of 60 min was considered optimal. In addition, Fig. 12 shows the XRD patterns of the calcines obtained at different roasting times, where it is apparent that the main calcine phases were fundamentally similar throughout,
Fig. 20. SEM-EDS pattern of the calcine. 7
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Fig. 21. SEM-EDS pattern of the leaching residue.
conditions, the resulting lixivium was found to contain ~3.43 g/L Li, ~25.50 g/L Na, ~11.00 g/L K, ~1.35 g/L Rb, and ~0.56 g/L Cs, as presented in Table 3. (See Figs. 15 and 16.) According to the optimal process conditions obtained in the previous experiments, the water-cum-material balance flowsheet was carried out on the process of leaching alkali metlas from lithium porcelain stone. The result is shown in Fig. 17.
NaCl. Meanwhile, calcium chloride facilitates this process. From Figs. 20 and 21, the soluble salt dissolves and the surface of leaching residue is sparse, porous, and dispersed after the calcine leached. This also indicates that the soluble salts of alkali metals were still leached, and then in the roasting process fully converted. After leaching, the contents of Na, K, S, Cl and the particle size of calcine clearly decreased, but the contents of O, Si, and Al did not leach during the water leaching process.
3.7. Analysis of calcine and residue
4. Conclusions
The roasting and leaching experiments were conducted under the optimum technological conditions of an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2, roasting temperature of 850 °C, roasting time of 60 min, solid/liquid ratio of 1/1, leaching at room temperature for 60 min, and stirring speed of 300 rpm. The calcine and leaching residue under the optimum technological conditions were analyzed by ICP-MS, XRD and SEM-EDS, respectively. The results are shown in Table 4 and Figs. 18~21. As shown in Table 4, the components of Li, Rb, Cs in the ore had been nearly complete leached and most of the F remained in residue. In addition, the leaching residue/lithium porcelain stone mass ratio was approximately 1.2. As shown in Fig. 18, new phases of NaCl, CaSO4, (Na,Ca)Al(Si,Al)3O8 were formed in the roasting process. Compared with Fig. 19, the most soluble salts were followed by water leaching. The extraction mechanism was found to be high-chemoselective ion-exchange between alkali metals in ore, and Na+ in Na2SO4/
We herein investigated the effects of the Na2SO4 and CaCl2 roasting additives on the selective water leaching of alkali metals from lithium porcelain stone. This procedure is expected to be a relatively cost-effective and straightforward method for the extraction of lithium, and an alternative to current industrial processes. Under the optimized roasting conditions of an ore/Na2SO4/CaCl2 mass ratio of 1/0.2/0.2, a temperature of 850 °C, and a time of 60 min, in addition to optimized leaching conditions of a liquid/solid ratio of 1:1, a temperature of 25 °C, a stirrer speed of 300 rpm, and a time of 60 min, the maximum leached contents of Li, Na, K, Rb, and Cs were 98.10, 49.80, 37.90, 97.15, and 98.24% respectively. Likewise, the concentrations of the alkali metals in the resulting lixivium were ~3.43 g/L Li, ~25.50 g/L Na, ~11.00 g/ L K, ~1.35 g/L Rb, and ~0.56 g/L Cs, thereby indicating the feasibility of this process. Furthermore, based on HSC modelling and X-ray 8
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diffraction measurements, the extraction mechanism was found to involve a highly chemoselective ion-exchange process between the alkali metals present in the ore, the Na+ present in Na2SO4, and the Ca2+ present in CaCl2. These results are of importance due to the anticipated increase in demand for Li, Rb, and Cs accompanied by an expected shortage in alkali metals available for extraction from ores.
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Declaration of Competing Interest We declare that we have no financial and personal relationships with other people or organizations that can inappropriately influence our work, there is no professional or other personal interest of any nature or kind in any product, service and/or company that could be construed as influencing the position presented in, or the review of, the manuscript entitled “Extraction of lithium, rubidium and cesium from lithium porcelain stone”. Acknowledgements This work was financially supported by the National Natural Science Foundation of China (No. 51564018 and No. 51764018), the Innovation Team Support Program of Jiangxi University of Science and Technology (3200826514), the Jiangxi Provincial Key Laboratory of Flash Green Developmet and Recycling (20193BCD40019), and the Program of Qingjiang Excellent Young Talents, Jiangxi University of Science and Technology. References Chen, Y., Tian, Q.Q., Chen, B.Z., Shi, X.C., Liao, T., 2011. Preparation of lithium carbonate from spodumene by a sodium carbonate autoclave process. Hydrometallurgy 109 (1), 43–46. He, M.S., Huang, X., Zou, G.S., 2014. Characteristics, Utilization Status and Suggestions of Lithium Containing Porcelain Stone (Soil) Mineral Resources in Jiangxi Province.
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