Preparation of calcium stannate from lead refining slag by alkaline leaching-purification-causticization process

Accepted Manuscript Preparation of calcium stannate from lead refining slag by alkaline leachingpurification-causticization process Wei Liu, Wenhua Li, Junwei Han, Dixiu Wu, Zihan Li, Kunhong Gu, Wenqing Qin PII: DOI: Reference:

S1383-5866(18)33548-2 https://doi.org/10.1016/j.seppur.2018.11.024 SEPPUR 15074

To appear in:

Separation and Purification Technology

Received Date: Revised Date: Accepted Date:

9 October 2018 26 October 2018 6 November 2018

Please cite this article as: W. Liu, W. Li, J. Han, D. Wu, Z. Li, K. Gu, W. Qin, Preparation of calcium stannate from lead refining slag by alkaline leaching-purification-causticization process, Separation and Purification Technology (2018), doi: https://doi.org/10.1016/j.seppur.2018.11.024

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Preparation of calcium stannate from lead refining slag by alkaline leaching-purification-causticization process Wei Liu1,2, Wenhua Li1, Junwei Han1,2*, Dixiu Wu1, Zihan Li1, Kunhong Gu1, Wenqing Qin1,2

1

School of Minerals Processing and Bioengineering, Central South University,

Changsha 410083, China. 2

Key Laboratory of Hunan Province for Clean and Efficient Utilization of Strategic

Calcium-containing Mineral Resources, Central South University, Changsha 410083, China. *Corresponding author: [email protected]

1

ABSTRACT: In the present study, a causticization method combined with alkaline leaching and purification process was used to treat with lead refining slag. In the alkaline leaching process, the effects of alkaline concentration, temperature, liquid to solid ratio, leaching time, and agitating speed were investigated, and the optimized leaching conditions were established: 2 mol/L of NaOH concentration, 2 h of leaching time, 85°C of temperature, 6 mL/g of liquid-to-solid ratio, 800 r/min of agitating speed. Under these conditions, the leaching rate of Sn reached 92% while the leaching rate of Pb was about 29%. After a purification process, the concentration of lead in the leachate was reduced to 0.0001g/L. Finally, high-purity calcium stannate (CaSnO3·3H2O) was obtained from the purified solution by the causticization process under the conditions: 5g of CaO dosage (per 200mL solution), 85°C of temperature, 2h of time and 800r/min of agitating speed. The precipitation rate of Sn was as high as 99.22%. Keywords: Alkaline leaching; Causticization; Lead refining slag; Purification; Tin

2

1. Introduction Lead acid batteries are widely applied in various vehicles and their grids generally contain plenty of tin-lead alloy, thus the used lead acid batteries will be a significant secondary resource for recycling lead. Pyrometallurgical process is the most widely used regeneration method of used lead acid batteries [1-4]. In the process, the used batteries were first smelted to form a crude lead with impurities such as Sn, Sb, and Bi [5-8]. It is necessary to remove these impurities from crude lead for obtaining a high-grade lead. The most common purification method is to add NaOH, NaNO3 into fused crude lead liquid, resulting in that tin and antimony are removed in the form of sodium stannate and sodium antimonite, respectively [9-11]. Meanwhile, sodium stannate, sodium antimonate, incomplete reaction of sodium hydroxide and a small amount of lead are separated to form lead refining slag [12, 13], which causes that the amount of lead refining slag gains a rapid growth with the development of secondary lead industry [14, 15]. Thus, lead refining slag contains many valuable metals and its improper management will also lead to heavy metal pollution. It is therefore urgent to develop an economical green technology to utilize the secondary resources. To effectively recover valuable metals from the slag, a number of methods have been developed for the treatment of lead refining slag, which can be summarized as follows: direct carbon reduction [16, 17] and cascade utilization [18]. Although many achievements in this field have been made, cascade utilization technologies has still not been widely applied in industry and direct carbon reduction technology is a 3

relatively better choice nowadays. The method of direct carbon reduction usually adopts a reduction smelting by reverberatory furnace. Lead, tin and antimony are readily to form alloy in the process and the residual alkali can play a role in maintaining the fluidity of slag system [5, 19-21]. However, there are three main problems presented in the conventional method. One of the problems is that the slag contains a lot of sodium hydroxide, which will react with the refractory of reverberatory furnace in the smelting process and thus lead to refractory corrosion [22, 23]. The second problem is that sodium stannate is difficult to be completely reduced in the reduction process, resulting in a low recovery of tin and the reduction slag containing high tin [24]. The last is that the resulting product is a Pb-Sn-Sb alloy, which needs further separation [25, 26]. Besides the separation of Pb-Sn-Sb alloy is very difficult due to the strong similarities among these three elements [27, 28]. As mentioned above, the tin contained in lead refining slag cannot be recycled effectively, which leads to an economic loss and bad battery performance when the secondary lead is used in battery production [29]. Therefore, it is necessary to explore an economical method to treat the slag, especially for the materials containing high tin. In the process of some solid wastes containing tin, such as waste printed circuit board and tin anode slime, hydrometallurgy method is cleaner and more efficient, and has been extensively applied [30-32]. Furthermore, it is found that CaSnO3 is easy to be reduced when the temperature is above 700°C according to thermodynamic calculation. Inspired by previous studies, a novel technology consisting of alkaline leaching, 4

purification, and causticization processes was developed for the extraction of tin and preparation of calcium stannate from lead refining slag. The effects of NaOH concentration, liquid-to-solid ratio, agitating speed, leaching temperature and time on the extraction of tin from the slag were detailly investigated based on thermodynamic calculations. The effects of CaO dosage, agitating speed, causticization temperature and time on the precipitation rate of Sn from the leachate after the purification process were also studied. Finally, the calcium stannate obtained were characterized by X-ray powder diffraction (XRD), thermogravimetric-differential thermo analysis (TG-DTA), and scanning electron microscopy combined with energy dispersive spectrometry (SEM-EDS).

2. Experimental 2.1 Materials The lead refining slag used in this study was obtained from a pyrometallurgical smelter of waste lead acid batteries in China. Firstly, the sample were dried at 100°C for 24 h and then were digested by HNO3-HCl system for chemical composition analysis by Inductively Coupled Plasma Optical Emission Spectrometer (ICP-OES; Thermo Scientific; iCAP 6500; Massachusetts; America). The results are presented in Table 1. It can be seen from Table 1 that the contents of tin and lead are 33.91% and 6.49%, respectively. Then the sample were further characterized by X-ray diffraction (XRD; D/MAX 2500; Rigaku, Almelo, Holland), and the results are shown in Fig. 1. The XRD pattern indicates that tin was in the form of mainly Na2SnO3 and some Sn,

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while lead was mainly in the form of Pb. 2.2 Experimental procedure The leaching experiments were conducted in an atmospheric pressure agitating leaching setup, whose schematic was given in our previous reports[33]. It was equipped with temperature controller, beaker, mechanical stirrer and water bath. For each test, NaOH solution with required concentration and volume was first prepared and placed in the beaker. The sample of 50 g was put into NaOH solution and leached at a given temperature. At the end of each experiment, leaching residue was dried, weighed, and ground followed by digestion for chemical composition analysis. Filtrate was measured for volume and sampled for chemical composition by ICP-OES. The filtrate was subjected to a purification process in a 400 mL glass beaker for removal of Pb by adding Na2S. Subsequently, the solution after purification was used for causticization experiment in the same equipment of leaching experimental setup. Besides, the solution after causticization process would be returned to the leaching process. A flowsheet of the whole process is presented in Fig. 2.

3. Results and discussion 3.1 E-pH diagrams Metal-H2O systems were used as a guide in pH and potential values of the leaching situation[30]. The calculation method of the E-pH diagram is as follows :(i) Determining the possible reactions in the system and the balance equations; (ii) Obtaining the standard electrode potential and deriving the equilibrium equation of E 6

and pH by calculating the Gibbs free energy with using the known thermodynamic data. Fig. 3 shows the E-pH diagrams of Sn-H2O and Pb-H2O system at 25°C. Note that the concentrations of related metallic ions are fixed at 1 mol/L, and both of the partial pressure of oxygen and hydrogen are at the standard atmospheric pressure of 101,325 Pa. The E-pH diagram of Sn-H2O system shown in Fig. 3(a) indicates that the transformation of Sn to Sn (Ⅳ) can be achieved in the whole pH range investigated. In alkaline solution, Sn (Ⅳ) exists as stannic ion, and it tends to precipitate as Sn(OH)4 and further dissolved as Sn4+ with the decrease in pH. As seen from Fig. 3(b), lead can exist in the form of PbO32

when the pH value is above 13,

indicating that the metallic lead in the slag can be oxidized and thus dissolved into the solution. 3.2 Alkaline leaching 3.2.1 Effect of NaOH concentration The concentration of NaOH plays an important role in a leaching process [34] , and therefore, the effect of NaOH concentration on the extraction of Sn was investigated. Other conditions were fixed as: 85°C of leaching temperature, 6 mL/g of liquid-to-solid ratio, 2 h of leaching time, 800 r/min of agitating speed. The results are presented in Fig. 4. It is seen from Fig. 4 that more than 70% Sn and 10% Pb were extracted from the slag when NaOH concentration was 0.5 mol/L. The leaching rate of Sn significantly increased as NaOH concentration was increased from 0.5 mol/L to 2.0 mol/L. Thereafter, it tended to keep constant when the concentration was higher than 2.0 mol/L. On the other hand, the leaching rate of Pb kept growing from 10% to 7

30% with increasing the concentration from 0.5mol/L to 2.5mol/L. As a result, NaOH concentration has a significant effect on the leaching of Sn and Pb from lead refining slag. NaOH concentration was therefore chosen as 2.0mol/L for recycling tin as more as possible, and all further experiments were carried out at this concentration. 3.2.2 Effect of leaching temperature It is well-known that leaching temperature is an important factor on a leaching process. Increasing temperature can accelerate reaction rate but consume more energy. To investigate the effect of leaching temperature on the extraction of Sn and Pb, a series of experiments were conducted from 25°C to 95°C, and the results are shown in Fig. 5. It can be seen from Fig. 5 that Sn extraction increased with the increase in temperature from 25°C to 85°C. Above 85°C, the leaching rate of Sn had no significant variation as the temperature increased. This indicated that increasing temperature within a certain range can contribute to the extraction of Sn. On the other hand, the leaching rate of Pb increased slightly with the increase in temperature from 25°C to 65°C, but it increased significantly when the temperature increased from 65°C to 95°C. Considering both energy consumption and evaporation of water at high temperatures, the optimal temperature was determined to be 85°C, at which the leaching rate of Sn and Pb were 92.71% and 28.76%, respectively. All further experiments were therefore carried out at 85°C. 3.2.3 Effect of liquid to solid ratio Liquid-to-solid ratio (L/S) is another important factor on the extraction of Sn and

8

Pb [35]. The effect of L/S on the extraction of Sn and Pb was therefore investigated and the results are presented in Fig. 6. As shown in this figure, the leaching rate of Sn increased slowly as the L/S increased from 5 to 6 mL/g and then remained unchanged when the L/S is above 6 mL/g. On the contrary, the leaching rate of Pb was stable when the L/S is less than 8 mL/g and then increased sharply as the L/S was above 8 mL/g. The result indicated that L/S had a more significant effect on Pb, and the optimal L/S was considered as 8 mL/g. 3.2.4 Effect of leaching time Fig. 7 shows that the effect of leaching time on the extractions of Sn and Pb. It can be seen from Fig. 7 that with the extension in leaching time, the leaching rate of Sn increased slowly as the leaching time increased from 0.5 to 2.0h and then remained unchanged when the leaching time is more than 2h. On the other hand, the leaching rate of Pb increased slowly when the leaching time is less than 2h and then had a relatively sharply growth when the leaching time was above 2h. The result indicated that increasing leaching time can promote the extraction of Sn and Pb. As a result, the optimal leaching time was considered as 2h. 3.2.5 Effect of agitating speed Agitating is also a key factor by reducing the boundary layer diffusion resistance and increase the diffusion rate of reactant on the extractions of Sn and Pb in a hydrometallurgical process [36]. The effect of agitating speed on the extraction of Sn and Pb was investigated, and the results are shown in Fig. 8. It was found that both the

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extraction of Sn and Pb increased with increasing agitating speed from 400r/min to 800r/min. When the speed is more than 800r/min, both the extraction of Sn and Pb change smoothly. As a result, the optimal agitating speed was determined as 800r/min. According to the above studies, the optimum conditions for the alkaline leaching experiment were established as: 85°C of leaching temperature, 2.0 mol/L of NaOH concentration, 8mL/g of liquid-to-solid ratio, 2h of leaching time, and 800r/min of agitating speed. In order to confirm the experimental results, two confirmation experiments were preformed, and the results are listed in Table 2. The results of confirmation experiments are consistent with that of the condition experiment, and more than 92% of Sn and about 29% of Pb can be extracted from the slag, while the leaching rates of other metals are relatively low. 3.3 Separation of tin and lead Based on the different sulfur ion affinity for metal ions in water, a purification process by adding Na2S was performed to separate lead from the leach solution. The purification experiments were carried out with 0.5mol/L Na2S solution at 85°C. The experimental results are shown in Table 3. It is seen that more than 99% of Pb was removed from the leach solution, and the precipitation rate of Sn could be limited below 5%. After the purification, the leach solution contains about 37.25 g/L Sn and less than 0.0001 g/L Pb, indicating that the purification process is very effectively.

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3.4 Causticization process 3.4.1 Effect of CaO dosage A causticization process with CaO powder was performed to obtain calcium stannate product from the leach solution after the purification process. The possible chemical reaction during the causticization process is as follows:

Na 2SnO 3 + CaO + H2O = CaSnO 3 ↓+2NaOH

(1)

It is well-known that CaO dosage plays a crucial role in the causticization process. Hence, the effect of CaO dosage was first investigated under the following conditions: 85°C of temperature, 2h of causticization time, and 600r/min of agitating speed. The results are presented in Fig. 9. It is seen from Fig. 9 that the precipitation rate of Sn increased significantly when CaO dosage was increased from 2g to 5g (per 200 mL the leach solution). More than 98% of Sn were precipitated when CaO dosage was 5g. Thereafter, with further increasing CaO dosage, the precipitation rate of Sn had no significant variation, indicating the causticization reaction had completed. Therefore, the optimal CaO dosage was considered to be 5 g per 200 mL the leach solution. 3.4.2 Effect of causticization temperature Temperature also plays an important role in a chemical reaction, and the effect of causticization temperature on the precipitation rate of Sn was investigated under the following conditions: 5g of CaO dosage, 2h of causticization time, and 600r/min of agitating speed. The results are presented in Fig. 10. It is seen from Fig. 10 that the precipitation rate significantly increases as the temperature increased from 25°C to

11

40°C. Thereafter, with further increasing the temperature, the value gradually increases and reaches its maximum (about 98%) at 85°C. Therefore, the optimum temperature was considered as 85°C. 3.4.3 Effect of causticization time Reaction time is another important factor on the causticization process and thus the effect of time on the precipitation rate of Sn was investigated under the following conditions: 5g of CaO dosage, 85°C of temperature, and 600r/min of agitating speed. The results are shown in Fig. 11. It is found that causticization time has a significant influence on the precipitation rate of Sn. When the time increased from 0.5h to 2h, the precipitation rate gradually increased from 45% to above 99%. Thereafter, the value had no significant change as the time was further increased. As a result, the optimum time was determined as 2h. 3.4.4 Effect of agitating speed Fig. 12 shows that the effect of agitating speed on the precipitation of Sn. It can be seen from Fig. 12 that the precipitation rate gradually increased from 35% to 99% as the agitating speed was increased from 200r/min to 600r/min, after which the value had no increase and even started to decrease with further increasing the agitating speed. Hence, the optimum agitating speed was considered to be 800r/min. Based on the above study, the optimum causticization conditions were established as: 85°C of temperature, 5g of CaO dosage per 200mL leach solution, 2h of time, and 800r/min of the stirring speed. Under these conditions, the precipitation rate of Sn

12

could reach 99%. In order to confirm the result, an expand experiment was conducted in a baker of 1L, and the results are listed in Table 4. As shown in Table 4, the Solution Ⅰ is the leachate after the purification process, the Solution Ⅱ is the leachate after the purification and causticization processes, and the Product is the white precipitation obtained from causticization process. It is seen from Table 4 that the concentration of Sn in the leach solution decreased from 37.25g/L to 0.29g/L after the causticization process, and the calcium stannate product contained 51.58% Sn. It is deduced that the precipitation rate of Sn was as high as 99.22%. Fig. 13 presents XRD pattern of the calcium stannate product obtained by the process. The results indicated that the solid product was calcium stannate, indicating the causticization was an effective method for the preparation of calcium stannate from the leach solution after a purification process. To further confirm the resulting product, the sample obtained was subjected to SEM-EDS and TG-DTA analyses, results of which are presented in Fig. 14 and Fig. 15, respectively. The SEM images (Fig. 14a and b) show that almost all the calcium stannate particles are smaller than 10µm. Fig. 14c and d indicate that the solid product is mainly composed of Ca, Sn and O, confirming that the sample is calcium stannate with high purity. It is seen from the TG-DTA curves (Fig. 15) that when the temperature is 451.3°C the mass loss rate is about 20%, which is consistent with the weight of crystal water in the molecule of CaSnO3·3H2O, demonstrating that the resulting product obtained by the process is mainly in the form of CaSnO3·3H2O.

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4. Conclusions Thermodynamic calculation and experimental results indicated that the tin contained in lead refining slag could be effectively extracted by alkaline leaching. The leaching parameters including NaOH concentration, temperature, leaching time, agitating speed and liquid-to-solid ratio have significant effects on the extraction of Sn. The optimized leaching conditions have been established as: 85°C of leaching temperature, 2.0mol/L of NaOH concentration, 8mL/g of liquid-to-solid ratio, 2h of leaching time, and 800r/min of agitating speed. Under these conditions, the leaching rate of Sn reached 92.98% while that of Pb was 28.52%. The lead dissolved in the leachate could be effectively removed by a purification process with Na2S solution, after which the concentration of Pb in the solution was as low as 0.0001g/L and the loss of tin is less than 5%. The purified solution was then subjected to a causticization process. Causticization temperature, time, CaO dosage and agitating speed have effects on the precipitation of Sn. More than 99% Sn could be precipitated after the process under the conditions: 5g CaO per 200mL solution, 85°C of temperature, 2h of time and 800r/min of agitating speed. Meanwhile, a high-purity calcium stannate (CaSnO3·3H2O) was obtained by the proposed process.

Acknowledgements The authors gratefully acknowledge the financial support by the National Natural Science Foundation of China (No. 51804342, No. 51874356, and No. 51604302), the Provincial Science and Technology Leader (Innovation Team of Interface Chemistry

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of Efficient and Clean Utilization of Complex Mineral Resources, No. 2016RS2016), the Collaborative Innovation Center for Clean and Efficient Utilization of Strategic Metal Mineral Resources, the Innovation Driven Plan of Central South University (Grant No. 2015CX005), Key Laboratory of Hunan Province for Clean and Efficient Utilization of Strategic Calcium-containing Mineral Resources (No. 2018TP1002), and the Scientific Research Starting Foundation of Central South University (No. 218041).

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Fig. 1 XRD pattern of the lead refining slag

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Fig. 2 The flowsheet of the proposed process for the treatment of lead refining slag

21

Fig. 3 E-pH diagrams of Metal-H2O systems at 25°C: (a) Sn-H2O system; (b) Pb-H2O system.

22

100

Extraction (%)

80

60

40

20

0 0.0

Sn Pb 0.5

1.0

1.5

2.0

2.5

3.0

NaOH concentration (mol/L) Fig. 4 Effect of NaOH concentration on the extraction of Sn and Pb (L/S=6mL/g, 85°C, 2h, 800r/min).

23

Fig. 5 Effect of leaching temperature on the extraction of Sn and Pb (L/S=6mL/g, [NaOH] = 2.0mol/L, 2h, 800r/min).

24

Fig. 6 Effect of leaching temperature on the extraction of Sn and Pb (85°C, [NaOH] = 2.0mol/L, 2h, 800r/min).

25

Fig. 7 Effect of leaching temperature on the extraction of Sn and Pb (85°C, [NaOH] = 2.0mol/L, L/S=8mL/g, 800r/min).

26

Fig. 8 Effect of agitating speed on the extraction of Sn and Pb (85°C, [NaOH] = 2.0mol/L, L/S=8mL/g, 2h).

27

Fig. 9 Effect of CaO dosage on the precipitation rate of Sn

28

Fig. 10 Effect of causticization temperature on the precipitation rate of Sn

29

Fig. 11 Effect of causticization time on the precipitation rate of Sn

30

Fig. 12 Effect of agitating speed on the precipitation rate of Sn

31

Fig. 13 XRD pattern of the resulting product

32

Fig. 14 SEM images of the product at 3000 times magnification (a), 5000 times magnification (b), BSE image of the product (c) and EDS pattern at point A (d).

33

Fig. 15 TG-DTA result of the calcium stannate product

34

Highlights:

High-purity calcium stannate (CaSnO3·3H2O) was obtained by the proposed process. Tin contained in lead refining slag was effectively extracted by alkaline leaching. More than 99% Sn could be precipitated from the solution after the causticization. Lead in the solution was effectively removed by a purification process with Na2S.

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Table 1 Main chemical composition of the lead refining slag (wt.%) Elements

Pb

Sn

Sb

Cu

S

Content

6.49

33.91

1.21

0.04

0.08

36

Table 2 Results of confirmation experiments under optimum condition (%) No.

Sn

Pb

Sb

Cu

Ⅰ

92.93

28.93

0.42

1.22

Ⅱ

93.02

28.11

0.48

1.02

Mean

92.98

28.52

0.45

1.12

37

Table 3 The results of the purification process Concentration (g/L) Name

V/mL Sn

Pb

Initial solution

200

38.99

2.27

After purification

200

37.25

<0.0001

4.46%

>99%

Precipitation rate

38

Table 4 Results of confirmation experiments under the optimum conditions (%) No.

Volume

Quantity

Sn

Pb

Solution Ⅰ

1.0L

-

37.25g/L

<0.0001 g/L

Solution Ⅱ

1.0L

-

0.29g/L

<0.0001 g/L

Product

-

71.66g

51.58%

<0.01%

39