RECOVERY OF VALUABLE METALS FROM SPENT LITHIUM-ION BATTERIES BY SMELTING REDUCTION PROCESS BASED ON
MnO-SiO2-Al2O3 SLAG SYSTEM
Ren Guoxing, Xiao Songwen*,Xie Meiqiu, Pan Bing, Fan Youqi, Wang Fenggang, Xia Xing Changsha Research Institute of Mining & Metallurgy, Changsha 410012, Hunan, China Keywords: spent lithium-ions battery, smelting reduction, MnO-SiO2-Al2O3 slag, pyrolusite
Abstract
Plenty of valuable metals, such as cobalt, nickel, copper, manganese and lithium, are present in spent lithium-ion batteries. A novel smelting reduction process based on MnO-SiO2-Al2O3 slag system for spent lithium ion batteries is developed, using pyrolusite ore as the major flux. And Co-Ni-Cu-Fe alloy and manganese-rich slag contained lithium are obtained. The results show that it is reasonable to control MnO/SiO2 ratio in the range of 2.05-3.23 (w/w) and Al2O3 content in 19.23-26.32wt.%, while the MnO and Li2O contents in the manganese-rich slag can reach 47.03 wt.% and 2.63 wt.%, respectively. In the following leaching experiments of the manganese-rich slag by sulphuric acid solution, the recovery efficiency of manganese and lithium can reach up to 79.86% and 94.85%, respectively. Compared with the conventional hydro-pyrometallurgical process of spent lithium-ion batteries, the present can preferably recover Mn and Li besides Co, Ni and Cu.
Introduction
With rapid market expansion of consumer electronics and electric vehicles, the consumption of lithium-ion batteries (LIBs) has been increased tremendously in recent years. At the same time, quantities of waste LIBs are generating worldwide, which could bring serious environmental issue. Spent LIBs contain plenty of valuable metals, such as cobalt, nickel, copper, manganese and lithium, as well as toxic ingredients [1-4]. Once discarding without proper treatment, both the environmental contamination and the waste of resource are inevitable. Therefore, developing advance technologies for spent LIBs recycling are of great significance.
The most common industrial processes for the recovery of valuable metals from LIBs are hydrometallurgy and pyrometallurgy methods[1]. Hydrometallurgical methods like Recupyl and Toxco processes are widely designed for specific type of batteries, which can’t adapt to the diversity of real spent LIBs [3]. In contrast, various types of used battery, such as LIBs with Al shells, polymer LIBs, and nickel metal hydride batteries, can be treated by a pyrometallurgical based process, such as Val’Eas and Inmetco processes [2]. Additionally, the pyrometallurgical based also has the advantage of higher throughput. A selection of suitable slag system is one of key factors that influence the efficiency of pyrometallurgical process. Now, the smelting
Advances in Molten Slags, Fluxes, and Salts: Proceedings of The 10th International Conference on Molten Slags, Fluxes and Salts (MOLTEN16) Edited by: Ramana G. Reddy, Pinakin Chaubal, P. Chris Pistorius, and Uday Pal TMS (The Minerals, Metals & Materials Society), 2016
reduction process of spent LIBs generally used the CaO-SiO2-Al2O3 slag or FeO-CaO-SiO2-Al2O3 slag system. However, the manganese and lithium can’t be recovered effectively by conventional hydro-pyrometallurgical process which is based on above slag systems [3, 4].
The present study proposes a new smelting reduction method of spent LIBs based on MnO-SiO2-Al2O3 slag system, in which the pyrolusite is used as the major slag former. After smelting, the Co, Ni and Cu from spent LIBs were concentrated into alloys, while the Mn and Li were enriched into slag phases. Effects of pyrolusite additon, SiO2 addition, CaO addition, and smelting temperature on the recoveries of Co, Ni and Cu were explored in detail. And the leachability of the manganese-rich slag obtained was also investigated.
Materials and Methods Materials
In this study, mixed spent LIBs which consist of LIBs with Al cans and polymer LIBs, were used as experimental materials. Table 1 presents the chemical composition of the LIBs. According to earlier study [3,4], the Al and C of spent LIBs exist as the form of metallic Al and graphite, respectively, which are used as reductant during the smelting process.
Table 1 chemical composition of LIBs used in this investigation.
Type Co Ni Cu Mn Fe Al Li C others
Al-cans 16.0 2.0 8.0 0.0 0.0 33.0 2.0 15.0 23.9 polymer 2.0 3.0 10.0 21.0 0.1 13.0 2.0 25.0 25.9
Pyrolusite is used as the major slag former. Table 2 shows the chemical composition of the pyrolusite. According to the XRD pattern of the pyrolusite given in Fig.1, it was found that the pyrolusite mainly consists of quartz (SiO2) and pyrolusite (MnO2).
Table 2 chemical composition of the pyrolusite in this study Elements MnO2 Fe P Al2O3 SiO2 CaO MgO contents/% 63.05 2.60 0.85 2.41 13.33 4.70 0.71
Fig. 1 XRD pattern of pyrolusite Experimental
To investigate the feasibility of the new process, the experiments were conducted using the procedure shown in Fig. 2. Smelting process consists of two steps: (1) firstly, the polymer LIBs were roasted at 800℃ for 2h in an muffle furnace to remove a partial carbon, and cooled in the furnace. The spent LIBs with Al cans were sheared into several small parts, and then mixed with the roasted polymer LIBs according to the mass ratios of LIBs with Al cans to the polymer LIBs=1:1. (2) Secondly, the mixed batteries, pyrolusite, and slag modifier (CaO and/or SiO2) were together put into an alumina crucible with a cover after being thoroughly mixed. They were then melted at pre-determine temperature for 30 min in an electrical melting furnace (Box Type) with MoSi2 alloy heating rods. After smelting, the specimens were cooled down naturally to ambient temperature (~25℃). The crucible was withdrawn from the furnace. Subsequently the alloy and slag were manually separated and ground for chemical analysis.
The manganese-rich slag obtained was firstly crushed, and ground to grain size <0.074 mm, then mixed with a determined amount of concentrated sulphuric acid (82 wt.%) corresponding to different acid to ore mass ratios. The obtained slurry was then heated at temperature 85℃ for 60min in a loft drier. After heating, the slurry was then leached with water at room temperature.
After 30 min the leached residues were filtered and dried for chemical analysis.
Polymer spent lithium-ion batteries(LIBs)
Pyrolusite Smelting reduction CaO+SiO2
Roasting Calcine LIBs with Al shells Mixing
Co-Ni-Cu-Fe alloy Co/Ni/Cu recovery
Rich-manganese slag Leaching
Solution Li/Mn recovery
Residue Filtration
Concentrated H2SO4
Fig. 2 Flow chart of Co, Ni, Cu, Mn and Li recovery from spent LIBs.
Results and Discussion Effect of parameters on the smelting reduction process
Effect of pyrolusite addition Fig. 3 shows that the effect of pyrolusite addition on the recovery of valuable metals. It can be seen that pyrolusite addition did not obviously effect on the recoveries of Co, Ni and Cu, but had a significantly influence on the Mn recovery in alloy. Mn recovery decreased from 19.36% to 1.31% as the pyrolusite addition increased from 1.5 (w/w) to 2.8 (w/w). A part of Mn in the batteries and pyrolusite was reduced into the alloy, which can be explained by the reduction of manganese oxide according to the following reactions [5,6]:
3(MnO)+2[Al]=3[Mn]+(Al2O3) (1)
(MnO)+C=[Mn]+CO（g） (2)
From an economic point of view, the lower composition of slag maker means less energy composition, so the pyrolusite addition of 1.5 (w/w) is appropriate.
1.4 1.6 1.8 2.0 2.2 2.4 2.6 2.8 3.0 3.2 0
10 20 30 40 50 60 70 80 90 100
Cu Co Ni Mn
recovery in alloy/%
pyrolusite addition （w/w）
Fig. 3 Effect of pyrolusite addition on the recovery of valuable metals (SiO2 addition 9.0%(w/w), CaO addition 2.0%(w/w), smelting temperature 1475℃, smelting time 30min) Effect of SiO2 addition Fig. 4 shows that the effect of SiO2 addition on the recovery of valuable metals. As verified in Fig. 4, the recoveries of Ni and Cu increased with the increasing of SiO2
addition from 0% (w/w) to 6.0% (w/w), and remained constant after that. Co recovery was about 99% for the studied conditions. Mn recovery in alloy slightly increased until the SiO2 addition below 4.0% (w/w). When the addition of SiO2 exceeded 4.0% (w/w), the Mn recovery decreased obviously. The reason was probably that the activity of MnO from slag decreased with the increase of the SiO2 addition, thus restraining the reductive ability of MnO and decreasing the manganese recovery [7]. The SiO2 addition of 6% (w/w) was appropriate.
-2 0 2 4 6 8 10 12 14 16
0 10 20 30 40 50 60 70 80 90 100
Mn Cu Co Ni
recovery in alloy/%
SiO2 addition/%
Fig. 4 Effect of SiO2 addition on the recovery of valuable metals (pyrolusite addition 2.0(w/w), CaO addition 2.0% (w/w), smelting temperature 1475℃, smelting time 30min) Effect of CaO addition Fig. 5 shows that the effect of CaO addition on the recovery of valuable metals. From Fig. 5, the recoveries of Co, Ni and Cu increased until CaO addition reached
2.0% (w/w), where 99.85% of Co, 99.52% of Ni, and 98.03% of Cu were recovered. Further increasing of CaO addition presented no benefit to metal recoveries. Mn recovery in alloy increased from 2.07% to 11.33% with the increasing of the CaO addition from 0% (w/w) to 8.0% (w/w). This is attributed to the increase of activity of MnO from slags according to previous study[8]. The appropriate CaO addition was 2.0% (w/w).
0 2 5 8
0 5 10 15 20 75 80 85 90 95 100
recovery in alloy/%
CaO addition/%
Cu Co Ni Mn
Fig. 5 Effect of CaO addition on the recovery of valuable metals (pyrolusite addition 2.0(w/w), SiO2 addition 9.0%(w/w), smelting temperature 1475℃, smelting time 30min)
Effect of smelting temperature Fig. 6 shows that the effect of smelting temperature on the recovery of valuable metals. It can be seen that the recoveries of valuable metals(Co, Ni and Cu) increased with the smelting temperature below 1475°C. When the smelting temperature exceeded 1475°C, the valuable metals recoveries kept nearly a constant. Therefore, the suitable smelting temperature was 1475°C.
1425 1450 1475 1500
0 10 20 80 90 100
smelting temperature/℃
Cu Co Ni Mn
recovery in alloy/%
Fig. 6 Effect of smelting temperature on the recovery of valuable metals (pyrolusite addition 2.0 (w/w), SiO2 addition 9.0% (w/w), CaO addition 2.0% (w/w), smelting time 30min) From the results mentioned above, the optimum condition of the smelting reduction experiment was pyrolusite addition of 1.5(w/w), SiO2 addition of 6.0%(w/w), CaO addition of 2.0% (w/w), smelting temperature 1475℃, and smelting time of 30min. The results of the optimum condition test are shown in Table 3. As shown in Table 3, the recovery rates of Co, Ni, and Cu were 99.79%, 99.30%, and 99.30%, respectively. The contents of MnO and Li2O in the slag obtained
were 47.03% and 2.63%, respectively. The major phases of the slag obtained were rich-calcium tephroite ((Mn,Ca)2SiO4) and galaxite (MnAl2O4) by the XRD analyze as shown in Fig. 7.
Table 3 experimental results under the optimum condition
item Cu Co Ni Fe Mn Li2O
Recovery in alloy/% 99.30 99.79 99.30 / 19.13 / Content in alloy/wt.% 21.78 24.43 5.05 10.85 30.13 / Content in slag/wt.% 0.11 0.02 0.02 0.19 47.03* 2.63
*- content of MnO in slag
Fig. 7 XRD pattern of manganese-rich slag under the optimum condition Table 4 results of slags vary different experimental conditions.
conditions contents in slag obtained/% Simplified slag MnO FeO Al2O3 SiO2 CaO MnO/SiO2 Al2O3
Pyrolysis addition
1.5 42.74 0.25 22.18 19.35 8.81 2.21 26.32 2.0 46.83 0.18 20.98 18.86 6.95 2.48 24.21 2.5 46.87 0.11 18.35 21.20 7.76 2.21 21.23
2.8 48.47 0.38 18.97 15.00 3.23 23.01
3.0 48.96 0.34 16.44 20.11 7.35 2.43 19.23 SiO2 addition
0.0 48.96 0.84 23.72 11.86 6.83 4.13 28.06 4.0 45.13 0.23 22.17 16.66 7.82 2.71 26.41
6.0 47.79 0.14 17.17 19.95 2.40 20.22
9.0 46.83 0.18 20.98 18.86 6.95 2.48 24.21 14.0 44.85 0.12 19.70 21.85 6.82 2.05 22.80 CaO addition
0.0 49.47 0.48 17.65 19.35 5.81 2.56 20.41 2.0 46.87 0.11 18.35 21.2 7.76 2.21 21.23 5.0 46.24 0.15 18.78 19.59 9.97 2.36 22.20 8.0 43.56 0.12 19.30 19.19 12.5 2.27 23.52 Analysis of appropriate slag composition
Table 4 summarizes the results of slag compositions at different experimental conditions. It can be seen that MnO, SiO2 and Al2O3 were the main compositions for obtained slags, so it is
feasible to describe the slag systems by the simplified ternary system MnO-SiO2-Al2O3. Haccuria et al has proposed that this system would be employed for the smelting process [9].
However, the suitable composition of slags wasn’t given. Assuming the compositions of the slag system is more suitable for the smelting reduction process of spent LIBs, when the recoveries of Co, Ni and Cu are up to 95%. Therefore, it can be concluded from Table 4 that the appropriate slag was MnO/SiO2=2.05-3.23 (w/w), and 19.23-26.32wt.% Al2O3 content. Moreover, It’s noted that the Al2O3 content in the manganese-rich slag is higher than that of slags for the Fe-Mn alloy production [10]. According to the results of Kim[11], the fluidity of these slags is very well under the temperature of 1475℃, so that the slag-alloy separation may be very easy.
Leachability of manganese-rich slag
Lithium from spent LIBs is collected into CaO-(FeO)-SiO2-Al2O3 slag during previous pyrometallurgical based process. However, the content of lithium in slags was less than 1.50 wt.%
due to the high addition of slag maker, which can’t be recovered by economical methods. By comparison, the lithium content in the manganese-rich slag was higher, up to 2.63 wt.% (see Table 3) in this study. This means that the extraction of lithium from the manganese-rich slag is more promising. In addition, most of manganese from spent LIBs and pyrolusite was enriched into the slag MnO-SiO2-Al2O3. Besides the Mn containing manganese-rich slag existed as the form of MnO due to the lower oxygen partial pressure of the smelting process according to previous study [12]. This means that MnO2 containing pyrolusite has been converted to acid-soluble MnO[13]. All of these are advantage to recover Li and Mn from the manganese-rich slag by the following leaching process.
In order to investigate the leachability of managanese-rich slag, further experiments were carried out under conditions of a sulfuric acid concentration of 82 wt. % with various acid to ore mass ratios, heating temperature of 85 °C, and heating time of 120 min. the results are depicted in Table 5. When the sulfuric acid addition exceeds 1.0 (w/w), the Mn and Li leaching efficiencies reached 79.86% and 94.85%, respectively. This means that the Li and Mn from spent LIBs can be effectively recovered by the new developed methods.
Table 5 results of the leachability of managanese-rich slag varying different sulfuric acid addition Sulfuric acid addition/ (w/w) Li leaching efficiency/% Mn leaching efficiency/%
0.75 80.13 65.66
1.00 94.85 79.86
1.25 94.14 79.72
1.50 92.41 76.39
Conclusions
The present study proposed a new smelting reduction method of spent lithium-ion batteries based on MnO-SiO2-Al2O3 slag system, in which the pyrolusite ore was used as the major slag former.
The Co-Ni-Cu-Fe alloy and lithium containing manganese-rich slag were obtained. The main results can be summarized as follows:
(1) Under conditions of the pyrolusite addition of 1.5 (w/w), SiO2 addition of 6%(w/w), CaO addition of 2% (w/w), smelting temperature 1475℃, and smelting time of 30min. 99.79% Co, 99.30% Ni, and 99.30% Cu were recovered and the contents of MnO and Li2O in the slag obtained were 47.03% and 2.63%, respectively. The major phases of the slag obtained were rich-calcium tephroite ((Mn,Ca)2SiO4) and galaxite (MnAl2O4).
(2) MnO-SiO2-Al2O3 slag system for the smelting process was appropriate under the condition of about MnO/SiO2=2.05-3.23 (w/w), and 19.23-26.32wt.% Al2O3 content.
(3) The lithium containing manganese-rich slag was leached in sulphuric acid media. The leaching efficiency of manganese and lithium reached 79.86% and 94.85%, respectively.
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