Green Process of Metal Recycling: Coprocessing Waste Printed

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A green process of metal recycling: Co-processing waste printed circuit boards and spent tin stripping solution Congren Yang, Jinhui Li, Quanyin Tan, Lili Liu, and Qingyin Dong ACS Sustainable Chem. Eng., Just Accepted Manuscript • DOI: 10.1021/ acssuschemeng.7b00245 • Publication Date (Web): 09 Mar 2017 Downloaded from http://pubs.acs.org on March 10, 2017

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ACS Sustainable Chemistry & Engineering

A green process of metal recycling: Co-processing waste printed circuit boards and spent tin stripping solution Congren Yang1, Jinhui Li1, *, Quanyin Tan1, Lili Liu2, Qingyin Dong2 1

State Key Joint Laboratory of Environment Simulation and Pollution Control, School of

Environment, Tsinghua University, Beijing 100084, China 2

Basel Convention Regional Centre for Asia and the Pacific, Beijing 100084, China

* Corresponding Author Address: Room 805, Sino-Italian Environmental and Energy-efficient Building, School of Environment, Tsinghua University, Haidian District, Beijing 100084, China E-mail address: [email protected] (J. Li) Tel.: +86-10-62794143, Fax: +86-10-62772048

ACS Paragon Plus Environment

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ABSTRACT Electronic waste (e-waste), including waste printed circuit boards (PCBs), has caused global concern owing to its potential environmental pollution and rich resource content. Previous studies have indicated that urban mining for metals recycling can decrease energy consumption and pollutants emission compared to the extraction of metals from natural minerals. During the production of PCBs, a large amount of spent tin stripping solution (TSS) is simultaneously generated, containing the significant amounts of metal ions and residue nitric acid. In this study, the co−processing of waste PCBs and spent TSS at room temperature was proposed and investigated, with the aim of developing an environmentally sound process to address these problems. This co−processing approach proved to be effective. 87% of the Sn−Pb solder, 30% of the Cu, 29% of the Fe, 78% of the Zn was leached from waste PCBs with spent TSS after 2 hours, at room temperature. Moreover, approximately 87% of the electronic components were dismantled from waste PCBs. About 99% of the Sn, Pb, Fe, Cu and Zn were recovered from the leaching solutions by chemical precipitation. The proposed green process has substantial advantages over traditional recovery methods of heating waste PCBs, in terms of both material and energy efficiency.

Keywords: E-waste; Waste printed circuit boards; Spent tin stripping solution; Green chemistry; Co-process; Metal recycling


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INTRODUCTION In order to achieve environmental and metal sustainability, green processes need to be developed for the recycling of metal from waste.1-4 For example, printed circuit boards (PCBs) are widely used in different fields such as electric and electronic equipment, information and sensing industries.

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Waste PCBs contain more than 40 metals,

including valuable metals (e.g. Sn, Fe, Cu, Zn) and hazardous substances (e.g. Pb, Cr, Cd).9, 10 The informal recycling of metals from waste PCBs, especially in developing countries, has caused serious environmental pollution and human health risks.11-15 Using the “Twelve Principles of Green Chemistry” is the most effective way to solve these issues while at the same time alleviating the shortage of mineral resources.16-19 Based on the principles of Green Chemistry, many environmentally sound processes have been developed to recover metal from e-waste. 20-25 Undoubtedly, the metals recycling from waste PCBs is necessary to the sustainable development of the electronics industry, and many techniques for waste PCBs dismantling and metals recycling have been developed. For dismantling the electronic components (ECs) and recovering the tin solder from waste PCBs, thermal treatments— such as infrared heater, electric heating tube, liquid-medium heating and solder-bath heating—are most commonly used.26-30 All these methods generally operate at about 250°C. For example, Zeng et al.31 used water-soluble ionic liquid as a heating medium to dismantle ECs and recover tin solder from waste PCBs, nearly 90% of the ECs were removed from the waste PCBs, at 250°C. Simultaneously, an automatic system for disassembling waste PCBs with heated air at 265 ± 5°C was developed by Wang et al.32. But more energy was consumed by heating waste PCBs above the melting point of tin


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solder. Recently, a chemical reagent was used to recover tin solder from waste PCBs. Yang et al. reported that 99% of tin could be leached with SnCl4 and HCl at 60 to 90°C, and the tin was then recovered from the purified solution by electro-deposition.33 HBF4 containing H2O2 was used to dissolve tin solder by Zhang et al., and almost 100% of the solder was dissolved.34 Recently, Zhang et al.

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reported that a leachant containing

methanesulfonic acid and hydrogen peroxide can also selectively dissolve the Sn-Pb solder. Although these techniques were conducted at lower temperatures, new chemical reagents (such as SnCl4, HCl, HBF4, H2O2) were consumed. For metals recycling from waste PCBs without ECs, electrostatic separation,36-39 wet jigging,40 froth flotation,40, 41 air current separation,42 etc., have been used to separate metals from nonmetals, pure metals can then be extracted from the mixed metals by vacuum metallurgy.43-46 Hydrometallurgy,47-51 bio-hydrometallurgy,52-56 and supercritical fluid,57-59 have also been applied, to leach valuable metals from waste PCBs. Valuable metals can be further recovered from leaching solution by adsorption-elution,60 electrowinning,61, 62 and so on. Spent tin stripping solution (TSS)—the tin, iron, copper and nitric acid containing waste solutions originally from PCBs production—is also classified as hazardous wastes.63 Nitric acid can be regenerated by solvent extraction-stripping64 and diffusion dialysis65,

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and the valuable metals can then be recovered by electrowinning,63,

66-68

precipitation,64-67 etc. The recycling of metals (e.g., Sn, Cu, et al.) is becoming more and more important, in order to counteract the depletion of mineral resources, especially as the demand for these metals continues to increase. While all of the methods discussed above have proved effective at recovering metals from waste PCBs, it is essential to develop


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more environmentally friendly processes that exhibit better performance in terms of both valuable-metal recycling and hazardous-substance control. In this study, a technique for co-processing waste PCBs and spent TSS at room temperature was developed and analyzed. This proposed process can not only meet these goals but also conserve energy, compared to traditional methods.

EXPERIMENTAL Characteristics of the samples Waste PCBs from desktop and laptop personal computers—after the central processing unit (CPU) and random access memory (RAM) had been removed—were used in all the experiments. First, ECs were removed from the PCBs by heating. The mass fractions of the ECs and the bare boards (without ECs) were 60% and 40%, respectively. The bare boards and ECs were crushed and screened to –1 mm. An appropriate amount of sample was dissolved in aqua regia for one day, and then the leached liquid was filtered through a 0.45 µm microfiltration membrane. The metal content in the filtrate was detected via inductively coupled plasma (ICP) (PE OPTIMA 8000). The results are presented in Table 1. Table 1 Composition of waste printed circuit boards (%) Element

Sn

Pb

Cu

Fe

Al

Zn

Ni

Cr

Bare boards 10.12 3.20 21.62 0.21 1.36 0.056 0.036 0.027 ECs

3.20 0.68 13.80 19.49 6.91 5.66 0.65 0.53

Total

6.0

1.7 16.9 11.8 4.7

3.4

0.4

0.3


Cd (g/t) Au(g/t) Ag(g/t) 0.53

/

194.91

14.45

40.76

112.68

8.9

24.4

145.7

5


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Most of the Sn in the spent TSS presented as insoluble hydrated stannic oxide. The spent TSS was therefore filtered through a 0.45 µm microfiltration membrane before determining the metal concentration via ICP. The results are presented in Table 2. Table 2 Metal concentrations in spent tin stripping solution Sn Cu Fe Pb Zn Ni Al Cr Cd H+ NO3− (g/L) (g/L) (g/L) (mg/L) (mg/L) (mg/L) (mg/L) (mg/L) (µg/L) (mol/L) (mol/L) 3.59 4.57 5.17 6.35 3.93 36.83 15.81 7.97 1.64 4.4 4.9

Leaching of waste PCBs with spent TSS The process discussed in this study is shown in Figure 1. At top of this figure is illustrated the method for co-processing the waste PCBs and the spent TSS. The leachate was then treated through five precipitation steps, with filtration in between each two steps. In order to the recovery Sn, Fe, Cu and Zn from the leaching solutions by chemical precipitation, the pH of the solution was adjusted to 1.5, 3, 6 and 8, respectively. A 98% H2SO4 was used to precipitate the Pb, the molar ratio of [SO42-]/[Pb2+] was always more than 1.4. All leaching experiments were conducted at room temperature. A piece of waste PCB without the CPU and RAM was placed into a plastic box (305 × 240 × 205 mm), and then 2 L of spent TSS was added to the box. The waste PCB was completely submerged within the spent TSS. Each 0.5 hour, 5 mL of solution was sampled to detect the target metals and H+ concentration. The soluble metals were determined with ICP. After the leaching, the PCB and ECs were washed with water and dried at room temperature, and the Sn, Pb, Fe, Cu and Zn were recovered from the leaching solution via chemical precipitation (Figure 1).


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Figure 1 Flow chart for dismanting and recoverying valuable metals from waste PCBs with spent TSS


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Small-scale pilot plant leaching experiments were also conducted at room temperature. 24.5 kg of waste PCBs without CPU and RAM were placed into a plastic basket (675 × 485 × 400 mm), and then the plastic basket was placed into a plastic box (755 × 530 × 500 mm). 194.5 kg of spent TSS was then added to the plastic box; the waste PCBs were completely submerged within the spent TSS. After leaching, the solution was sampled, to detect the target metals via ICP, and the PCBs and ECs were washed with water and dried at room temperature.

Recovery of valuable metals The composition of metals in solution changed with pH was calculated using Visual MINTEQ. All chemical reagents used in the experiments were analytical grade. First, 32% NaOH solution was added with a dropper, to 200 mL of the leaching solution, with a stirring speed of 160 rpm at room temperature for 1 hour, and the endpoint pH was adjusted to 1.5. The tin precipitate was filtered, washed with deionized water and dried at 60°C. 2.5 mL 98% H2SO4 was then added into the filtrate to precipitate the lead. The iron, copper and zinc precipitation was similar to the precipitation of tin, but the pH endpoints were 3, 6 and 8, respectively (Figure 1). The precipitate was examined using X-ray diffraction (XRD) (BRUKER D8 ADVANCE) and X-ray fluorescence (XRF) (Shimadzu XRF-1800). The pH value of the solution was measured with a pH meter (METTLER TOLEDO Five Easy).

RESULTS AND DISCUSSION Characteristics of waste PCBs leached with spent TSS


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The characteristics of waste PCBs leached with spent TSS, in terms of changes in the concentrations of H+, concentrations of dissolved tin ions and metal extraction percentage, are shown in Figure 2. The concentrations of H+ decreased from 4.4 mol/L to 2.2 mol/L in the first 2 hours (Figure 2A) due to the consumption of nitric acid. In the leaching process, a metal was oxidized to its ions by nitric acid; for instance, Sn0 was oxidized to Sn4+

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. Then, within the next 1 hour, the concentrations of H+ further

decreased (albeit more slowly) from 2.2 to 2.0 mol/L. The Sn and Pb in the waste PCBs are in the form of Sn−Pb solder. The Sn−Pb solder quickly was dissolved from waste PCBs. The Sn concentration increased from 3.6 g/L to 11.9 g/L, and the Pb extraction percentage reached 81% after 1 hour (Figure 2B). Over the subsequent 2 hours, the dissolution of Pb proceeded more and more slowly, and the Pb extraction percentage finally reached 90% after 3 hours. The Sn concentration in solution was 15.8 g/L after 3 hours. It is well known that tin reacts with nitric acid, which converts it into an insoluble hydrated stannic oxide.67, 69 When we used ICP to detect the metal concentrations in the leach liquid, the leach liquids had to be filtered through a 0.45 µm microfiltration membrane. Since Pb presents as Pb2+ ions in solution, the Pb extraction percentage can be used to indicate the dissolution of Sn−Pb solder. In other words, 90% of the Pb was extracted from the waste PCBs, indicating that 90% of the Sn−Pb solder was dissolved. The Cu extraction percentage increased with time, and after 3 hours 36% of the Cu had been leached (Figure 2C). Scott et al.67 reported that at higher nitric acid concentrations (> 2.1 mol/L), both Sn−Pb solder and Cu dissolved quickly, while at lower nitric acid concentrations (< 2.1 mol/L) the Sn−Pb solder still dissolved rapidly, but the dissolution rate of Cu proceeded much more slowly.


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B 25

100

Sn Pb

90

20

Sn concentration (g/L)

+ H Concentration (mol/L)

4

3

2

1

80 70

15

60 50

10

40 30

5

Pb extraction (%)

A5

20 10

0 0.0

0.5

1.0

1.5

2.0

2.5

3.0

0 0.0

3.5

0.5

1.0

Time (h)

C

90

80

1.6

80

70

1.4

70

60

1.2

50

1.0

40

0.8

30

0.6

20

0.4

20

10

0.2

10

0.5

1.0

1.5

2.0

2.5

Cu extraction (%)

Al extraction (%)

1.8

Cu Al

Metal extraction (%)

100

0 0.0

2.0

2.5

3.0

0 3.5

D

2.0

90

1.5

Time (h)

100

60 50 40 30

0 0.0

0.0 3.5

3.0

Fe

0.5

1.0

1.5

2.0

2.5

3.0

3.5

Time (h)

Time (h)

F

E

7

100

Ni Zn

90

6

Cr Cd

80 5

70



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60 50 40 30

4 3 2

20 1

10 0 0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

0 0.0

0.5

1.0

Time (h)

1.5

2.0

2.5

3.0

3.5

Time (h)

Figure 2 Concentrations of H+ in solution (A), concentrations of Sn in solution and Pb extraction (B), metal extraction (C-F) during the leaching of waste PCBs

The Fe and Zn extraction percentages increased with time, finally leaching out about 48% and 83% after 3 hours, respectively. Simultaneously, approximately 40% of the Ni


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was leached out after 3 hours. Only 0.2% of the Al was leached out with nitric acid after 3 hours owing to the formation of passivation film (Figure 2C-E). In addition, the dissolution of hazardous substances (such as Cr and Cd) was also investigated (Figure 2F). Unlike the other metals, the Cr concentration in the leaching solution increased from 7.97 to 12.02 mg/L, yet the extraction percentage of Cr was less than 0.6% after 3 hours. The Cd concentration in the leaching solution increased from 1.64 to 99.12 µg/L (an increase of about 60-fold), and the extraction percentage of Cd was about 5% after 3 hours. Table 1 shows most of the Cr and Cd present in the ECs. It can be concluded from the results that 88% of the Sn-Pb solder, 26% of the Cu, 44% of the Fe, 71% of the Zn, and 36% of the Ni, but only 0.2% of the Al, 0.5% of the Cr, and 5% of the Cd were leached after 2 hours. Meantime, most of electronic components had been removed from the waste PCBs after 2 hours. Based on these results, a time of 2 hours was selected for the subsequent experiments. The experiments were repeated, and the results are shown in Figure 3. Figure 3 shows that about 87% of the Sn-Pb solder, 30% of the Cu, 29% of the Fe, and 78% of the Zn was leached out after 2 hours. The dissolution of Sn-Pb solder and Cu was similar for each waste PCB, and the relative deviation was less than 5%. While the dissolution of Fe and Zn differed from one PCB sample to another, the relative deviations for Fe and Zn were about 26% and 16% respectively, because most of the Fe is present in the external interface and CPU socket, as is shown in Figure 4 in the red zone. Most of the Zn, Ni and Cr are present with Fe as anticorrosive coating and/or stainless steel. If the anticorrosive coating is intact, it is difficult to dissolve iron with nitric acid, whereas if the anticorrosive coating is destroyed, it is easy to dissolve the iron with nitric acid. As a


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result, the dissolutions of Fe and Zn were similar to each other, but could be very different for each waste PCB. Simultaneously, more than 87% of the ECs could be removed from waste PCBs (also seen in Figure 4). The proportion of ECs removed from waste PCBs (η) can be estimated by considering the weight. After 2 hours of leaching time, the removed ECs and PCBs were washed with tap water and dried naturally, and the weight of the removed ECs was recorded as W1. The ECs (such as chips) remaining on the PCBs were removed by heating, and that weight was recorded as W2. Therefore, the proportion of ECs removed from waste PCBs by leaching was calculated as follows:

η=

W1 × 100 W1 + W2

(1)

100 90 80


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70 60 50 40 30 20 10 0 Pb

Zn

Fe

Cu

Element

Figure 3 Extraction percentage of metal after 2 hours of leaching


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Figure 4 Waste PCBs before (on left) and after (on right) leaching with spent TSS

Figure 4 also shows that some ECs, such as chips, remained on the PCBs after 2 hours of leaching, because of variations in the EC packaging. Over the years, EC packaging has evolved from though hole packaging, to surface mount packaging, to area array packaging (Figure S1).31 The single in-line (SIP) or dual in-line (DIP) package can be mounted by through−hole soldering, and the ECs could be easily separated from waste PCBs when the solder was opened. The surface mount packaging was used to mount or place ECs directly onto the surface of the PCBs with solder, and when the solder was opened and/or the wire was dissolved, the ECs could be split off from the PCBs. The ball grid array (BGA) is one type of area array; it interconnected chips to the PCB with solder bumps that have been deposited onto the chip pads (Figure S1-2). After 2 hours of leaching, the outside solder bumps were dissolved with nitric acid, while tin reacted with nitric acid and was converted to an insoluble hydrated stannic oxide. The hydrated stannic oxide hinders the diffusion of the solution into the interior, so that the chips cannot be removed from the PCB because the inside solder bumps cannot be dissolved (Figure S2).


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Simulated small-scale pilot plant leaching of waste PCBs with spent TSS The small-scale pilot plant leaching experiments discussed here are described at the end of Section 2.2. The outcomes of these experiments are shown in Table S1. After 2 hours of leaching time, the Sn, Cu, Fe and Zn extraction percentages were 98.5%, 27.6%, 43.5% and 51.1%, respectively. At the same time, the Pb concentration in solution increased from 2.65 to 499.9 mg/L, and similarly the Al and Ni concentrations increased from 9.15 to 72.5 mg/L, and 6.25 to 256.9 mg/L, respectively. The dissolution of hazardous substances (such as As, Cr and Cd) was also investigated. Compared with the other metals leached with spent TSS, the As concentration in solution increased from 0 to 3.2 mg/L, the Cr concentration from 9.75 to 13.0 mg/L, and the Cd concentration from 1.0 to 37.52 µg/L. Most of ECs were removed from the waste PCBs, but a small amount of the ECs, such as chips, remained on the PCBs after 2 hours of leaching because of the special packaging (Figure S3). The results of the small-scale pilot plant leaching experiment were very close to those are mentioned in Section 3.1 for the lab experiments. Therefore, spent TSS can be used to dismantle waste PCBs and to leach Sn, Cu, and other valuable metals from waste PCBs.

Recovery of valuable metals from leaching liquids Simulated precipitation of valuable metals In this section, the recovery of Sn, Pb, Fe, Cu and Zn from leaching solutions is analyzed further. First, chemical forms of these metals in solution were calculated with Visual MINTEQ, and the results are shown in Figure 5.


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Figure 5A shows that hydrolysis of Sn4+ to Sn(OH)4 occurs very easily. More than 99% of the Sn4+ was precipitated as insoluble Sn(OH)4(s) at pH 1.5; and the insoluble Sn(OH)4(s) converted into an soluble Sn(OH)62− when the solution pH was more than 9. When the solution pH was less than 1.79, the hydrolysis of Fe3+ to FeOH2+ and Fe2(OH)24+ with increasing pH caused a decrease in the percentage of Fe3+ (Figure 5B). The soluble iron ions began to form Fe(OH)3(s) at pH 1.79, resulting in a decrease in the percentages of Fe3+, FeOH2+ and Fe2(OH)24+. More than 99% of the iron was precipitated at a pH of 2.7 (Figure 5B). More and more of the Cu2+ and CuNO3+ ions were converted into CuSO4(aq) because HSO4− lost a hydrogen ion and transformed into SO42− with increasing the pH. The precipitation of the soluble copper as Cu4SO4(OH)6•H2O(s) led to a decrease in the percentage of Cu2+, CuNO3+ and CuSO4(aq) when the solution pH was more than 4.32, and more than 99% of the copper was precipitated at a pH of 5.75. With a further increase of the pH to 8.70, the Cu4SO4(OH)6•H2O(s) began to convert into an insoluble Cu(OH)2(s) (Figure 5C). Similar to copper, Zn2+ and ZnNO3+ transformed into ZnSO4(aq) with increasing pH. The soluble zinc began to form Zn4SO4(OH)6(s) at pH 6.30, and about 99% of the zinc was precipitated at a pH of 7.65, while at the same time the Zn4SO4(OH)6(s) began to convert into an insoluble Zn(OH)2(s) at this pH value (Figure 5D). When concentration H2SO4 was continuously added into the solution, increasing the molar ratio of [SO42-]/[Pb2+], the percentages of Pb2+, PbNO3+ and Pb(NO3)2(aq) decreased because of the precipitation of soluble lead as PbSO4(s). And 99% of the Pb2+ was precipitated when the molar ratio of [SO42-]/[Pb2+] was more than 1.4 (Figure 5E).


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Meanwhile, sulfate, such as Fe2(SO4)3, CuSO4, ZnSO4, can be used to replace H2SO4 as a lead precipitant. This process, however, is not discussed further in this report.

A

B

100

100

Sn(OH)6

Sn(OH)4(s)

80 70 60 50 40 30 4+

Sn

20

Fe

90

Chemical forms of Fe (in %)

Chemical forms of Sn (in %)

3+

2-

90

10

Fe(OH)3(s)

80 70 60 50 40 4+

Fe2(OH)2

30 20

5+

Fe3(OH)4

2+

FeOH

10

0 0

1

2

3

4

5

6

7

8

9

10

11

12

13

0 1.0

14

1.5

2.0

2.5

pH

4.0

100

Cu(OH)2(s)

2+

70

Cu

60 50 40 +

30

CuNO3

20 10

80

2+

Zn 70 60 50 40 30 +

20

ZnNO3

0

1

2

3

4

5

6

7

8

9

10

11

12

13

14

Zn(OH)4

2-

Zn(SO4)2

0

0

ZnSO4(aq) 2-

10

CuSO4(aq)

Zn(OH)2(s)

Zn4SO4(OH)6(s)

90

Chemical forms of Zn (in %)

Cu4SO4(OH)6(H2O)(s)

80

Chemical forms of Cu (in %)

3.5

D

C 90

3.0

pH

100

0

1

2

3

pH

4

5

6

7

8

9

10

11

12

13

14

pH

E

100

PbSO4(s)

90

Chemical forms of Pb (in %)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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80 70 60

2+

Pb

50 +

PbNO3

40 30

Pb(NO3)2(aq)

20 10 0 0.0

0.5

1.0

1.5

2.0

2.5

3.0

3.5

4.0

4.5

5.0

22+ [SO4 ]/[Pb ]


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Figure 5 Chemical forms of the five metals in solution changed with the pH or molar ratio of [SO42-]/[Pb2+]. A−Sn4+ 0.1 mol/L, HNO3 0.5 mol/L, B−Fe(NO3)3 0.15 mol/L, HNO3 0.1 mol/L, C−Cu(NO3)2 0.2 mol/L, HNO3 0.1 mol/L, 0.1 mol/L H2SO4, D−Zn(NO3)2 0.2 mol/L, HNO3 0.1 mol/L, H2SO4 0.1 mol/L, E−Pb(NO3)2 0.02 mol/L, HNO3 0.1 mol/L As can be concluded from above, the pH values selected for precipitation of Sn, Fe, Cu and Zn were 1.5, 3, 6 and 8 respectively, and about 99% of the soluble metals were precipitated at these pH values. About 99% of the soluble lead was precipitated when the molar ratio of [SO42-]/[Pb2+] was more than 1.4.

Recovery of tin, lead and iron Figure 6A shows that 98.6% of the Sn was recovered by precipitating soluble tin as Sn(OH)4. This is an important result, but 92.2% of Fe was also co-precipitated as amorphous Fe(OH)3 because the initial pH for precipitation of Fe was 1.79, which is very close to the 1.5 for Sn. Simultaneously, 45.0% of the Cu and 43.9% of the Zn were also co-precipitated with Sn, although the initial pH values for precipitation of Cu and Zn were 4.32 and 6.30 respectively, which are much farther from the 1.5 value. Figure 6B shows that as a result the tin precipitate contained 54.60% Sn, together with 12.39% Fe and 2.72% Cu. The XRD pattern of tin precipitate is shown in Figure 7A, but only the low intensity characteristic peaks of SnO2 were detected by this method; in fact, the Sn(OH)4 was hydrated stannic oxide (SnO2•xH2O). The co-precipitation of Fe and Cu was also observed for the iron precipitation process at a pH of 3 (Figure 6). The precipitation of Cu and Zn with Sn could be attributed to the sorption of metals ions on


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amorphous Fe(OH)3.70, 71 Farley et al.70 proposed a surface precipitation model for the sorption of metal ions (such as Cu2+, Zn2+, Pb2+) on iron hydroxide. According to the model, Langmuir adsorption can be observed at low metal concentrations; as metal concentration is increased, Freundlich adsorption can be observed, where both adsorption and solid-solution precipitation occur simultaneously; and solid-solution precipitation become dominant over the sorption process when the metal concentration is further increased. Through the process described above 98.6% of the tin is captured. This fraction is a substantial addition in terms of amount to existing Sn upgrading processes. The Fe and Cu can also be further separated and recovered from the tin precipitate during smelting. The content and recovery of Pb in lead precipitate were 64.84% and 63.9%, respectively (Figure 6). The XRD patterns of lead precipitate are shown in Figure 7B, and the peaks were well in agreement with the characteristic peaks of PbSO4. The remaining Pb in solution was further precipitated as PbSO4 crystals with precipitating Fe3+ (Figure 6 and Figure 7C). An iron precipitate containing 39.37% Fe, 8.12% Pb and 12.69% Cu was obtained, and the recovery of Fe, Pb and Cu were 7.8%, 19.7% and 11.2%, respectively (Figure 6). The XRD patterns of iron precipitate are shown in Figure 7C, where the peaks are in agreement with the characteristic peaks of PbSO4, but only a low intensity peak of Fe(OH)3 was detected, at about 2-theta of 35°. During the lead precipitation process, excess sulfuric acid was added, and more than half of the SO42- in solution presented as HSO4− when the pH was less than 1.5. HSO4− lost a hydrogen ion and transformed into SO42− with increasing pH, and the percentage of HSO4− was less than 2% when pH


18

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reached 3. Thus, the remaining Pb2+ in solution further precipitated as insoluble PbSO4 crystals with increasing concentration of SO42−. In this way the remaining Pb and Fe in the solution were taken out. These metals can interfere with the upgrading process of Cu and Zn.

A 100

Tin precipitate

90

Lead precipitate

Recovery (%)

80 70

Iorn precipitate

60

Copper precipitate

50

Zinc precipitate

40 30 20 10 0

Sn

Pb

Fe

Cu

Zn

Elements

B 80 70

Weight percentage of element (%)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60


Sn

60

Pb

50

Fe

40

Cu

30

Zn

20

S Ni

10 0

Tin

Lead

Iron

Copper

Zinc

Precipitate

Figure 6 Elements recoveries (A) and weight percentages (B) in precipitate


19


A

B

500

2500

△ SnO2

Intensity (Counts)

Intensity (Counts)

2000

300 △ △

200

△

100

0 10

20

30

40

50

60

◆

70

80

◆

1500 ◆◆ ◆

◆

1000

◆ ◆ ◆ ◆

◆◆ ◆

◆

◆

◆ ◆

◆

◆ ◆

◆

◆

◆ ◆ ◆◆ ◆ ◆

◆

0

90

10

20

30

2-Theta (°)

40

50

60

70

80

90

2-Theta (°)

C

D

500

2500 ◆

PbSO4

◇ Cu4SO4(OH)6(H2O)

○ Fe(OH)3 ◆

400 ◆

2000

◆

300

◆ ◆ ◆ ◆◆

200

◇

◆

Intensity (Counts)

Intensity (Counts)

PbSO4

◆

500

△

◆

◆

◆

400

◆

◆

○

◆ ◆

◆ ◆

◆

100

1500

◇

1000

◇ ◇◇ ◇◇ ◇

◇◇

500

0

◇

◇ ◇

0 10

20

30

40

50

60

70

80

90

10

20

2-Theta (°)

30

40

50

60

70

80

90

2-Theta (°)

E

1400

▽

1200

Intensity (Counts)

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46 47 48 49 50 51 52 53 54 55 56 57 58 59 60

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▽▽

▽

▲

(Zn, Cu)7(SO4)2(OH)10•3H2O (Zn, Cu)4SO4(OH)6•4H2O

1000 ▽

800

▽

▲ ▽

▽ ▽

▽

600

▽

400 200 0 10

20

30

40

50

60

70

80

90

2-Theta (°)

Figure 7 XRD patterns of tin (A), lead (B), iron (C), copper (D) and zinc (E) precipitate

Recovery of copper and zinc As shown in Figure 6, a copper precipitate containing, on an element basis, 49.6% Cu and 14.50% Zn, was obtained, accompanying 43.4% Cu recovery and 22.1% Zn recovery. The XRD patterns of copper precipitate show that the peaks were in agreement


20

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with the characteristic peaks of Cu4SO4(OH)6•H2O (Figure 7D). A similar result was observed by Giannopoulou and Panias71. The content and recovery of Zn in zinc precipitate were 56.6% and 24.4%, respectively, while at the same time, 7.70% Ni was present in the zinc precipitate. The XRD patterns of zinc precipitate show that the main precipitate was (Zn, Cu)7(SO4)2(OH)10•3H2O, together with lesser amounts of (Zn, Cu)4SO4(OH)6•4H2O (Figure 7E). Generally speaking, Cu(OH)2 can be obtained by neutralization with NaOH at pH of 6, yet while, the XRD analysis shows in the present case that the precipitation of Cu4SO4(OH)6•H2O was obtained at a pH of 6, neither Cu(OH)2 nor Cu2NO3(OH)3 was obtained. In order to explain this result, solubility products of the various copper compounds involved were considered. Comparing the solubility products (Ksp) shown in Table

3,

the

order

for

Cu

precipitation

is

as

follows:

Cu4SO4(OH)6

>

Cu4SO4(OH)6•H2O > Cu2NO3(OH)3 > Cu(OH)2. It was concluded therefore that Cu4SO4(OH)6 and/or Cu4SO4(OH)6•H2O preferentially precipitate.72 The Cu4SO4(OH)6 or Cu4SO4(OH)6•H2O can be obtained by titration of CuSO4 solution with NaOH solution.73-76 Furthermore, the Cu4SO4(OH)6 was synthesized with urea and Cu(NO3)2 solutions containing SO42−, and the addition of sulfate caused precipitation to occur earlier.77 Cu4SO4(OH)6•H2O can also be obtained by using CuSO4 and urea as starting solutions; these can be transformed into Cu(OH)2 by adding NaOH into the solution.72 In spite of this possibility, only the characteristic peaks of Cu4SO4(OH)6•H2O were detected, in practice, with XRD (Figure 7D). The Cu4SO4(OH)6•H2O was considered as a type of metastable form of Cu4SO4(OH)6 by Zittlau et al.. The hydrate form was observed to precipitate between 20 and 40°C with pH values up to 10, but the Cu4SO4(OH)6•H2O will


21


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only transform into Cu4SO4(OH)6 at 40–50°C.75 This difference can explain why the Cu4SO4(OH)6•H2O precipitated first from the Cu(NO3)2 solution containing NO3− and SO42−, although the concentration of NO3− was far higher than that of SO42−. Table 3 Solubility products of the various copper compounds Reaction

Ksp

Cu(OH)2 (s) → Cu 2+ (aq)+2OH − (aq)

2.2 × 10−20

78

(2)

Cu 2 NO3 (OH)3 (s) → 2Cu 2+ (aq)+NO3− (aq)+3OH − (aq)

1.9 × 10−36

79

(3)

Cu 4SO4 (OH)6 (s) → 4Cu 2+ (aq)+SO42− (aq)+6OH − (aq)

1.01 × 10−69 74

(4)

1.58 × 10−70 75

Cu 4SO4 (OH)6 H 2O(s) → 4Cu 2+ (aq)+SO42− (aq)+6OH − (aq)+H 2O *

3.29 × 10−67 *

(5)

The Ksp of Cu4SO4(OH)6•H2O was calculated to be 3.29 × 10−67 based on Zittlau et

al.75

Alternatively, the formation of Cu4SO4(OH)6•H2O may be due to anion-exchange reactions of Cu2NO3(OH)3. Cu2NO3(OH)3 is a layered structure compound [like Cu2(OH)4], in which one of the OH− ions in Cu2(OH)4 has been substituted by NO3−. This NO3− ion is incorporated in the interlayer between the Cu2(OH)3 layers.80, 81 Furthermore, many research reports have indicated that the NO3− ions in (M, Me)2NO3(OH)3 compounds [such as Cu2NO3(OH)3, (Cu, Zn)2NO3(OH)3, (Cu, Ni)2NO3(OH)3, etc.] can be replaced by monovalent anions (such as OH−, Cl−). Also, divalent anions (such as CO32−, SO42−), organic anions (such as acetate, terephthalate, benzoate, alkyl sulfate and alkanesulfonate ions) can replace NO3−.80-84 For instance, Meyn et al. reported that the NO3− ions in Cu2NO3(OH)3 were very easily replaced by alkyl sulfate and


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alkanesulfonate ions.80 Newman et al. prepared Cu2(Ac)(OH)3 by exchanging NO3− with acetate anions (Ac).81 Jimenezlopez et al. found that the Ac in Cu2(Ac)(OH)3 could be substituted by I−, simultaneously as a that anion-exchange reactions Ac−/I− was reversible.84 Stanimirova et al. stated that the NO3− in Cu2NO3(OH)3 can be replaced by Cl− to form Cu2Cl(OH)3. In this case, the Cu2Cl(OH)3 transformed into Cu4SO4(OH)6 by exchanging Cl− with SO42−.85 Figure 5D shows that Cu4SO4(OH)6•H2O(s) is converted into Cu(OH)2 when the solution pH is more than 8.70. Similar results were observed by Kratohvil et al.72 The Zn precipitation from leaching solution is similar to that for Cu, and hence we need not discuss the Zn case here in detail again. Especially, except the anion-exchange reactions of basic Zn salts [such as Zn5(OH)8(NO3)2•2H2O],80, 81 some of the Zn can also be substituted by other metal ions to form hydroxy double salts [such as (Zn, Me)5(OH)8(NO3)2•2H2O], where the ionic radii difference between Zn and Me should be not more than 0.05 Å, such as is the case with Zn2+ and Co2+, Zn2+ and Ni2+, and Zn2+ and Cu2+.80

CONCLUSIONS Environmentally friendly and efficient technologies for recycling of metals from waste PCBs need to be developed. This study represents an important contribution towards this goal. A green co-processing technology is here proposed, which can be applied to the recovery of valuable metals from waste PCBs and spent TSS at room temperature.


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The proposed green process proved to be effective for waste PCBs dismantling and metals recycling. Approximately 87% of the Sn-Pb solder was successfully leached from waste PCBs with spent TSS after 2 hours. Simultaneously, more than 87% of the ECs were removed from waste PCBs, and 30% of the Cu, 29% of the Fe and 78% of the Zn were also leached from waste PCBs. Small-scale pilot plant leaching experiments were conducted as well, and similarly effective results were obtained. Based on the solution composition calculation, the pH values selected for the precipitation of Sn, Fe, Cu and Zn were 1.5, 3, 6 and 8 respectively, and the molar ratio of [SO42-]/[Pb2+] selected for the precipitation of Pb should be more than 1.4. About 99% of the Sn, Pb, Fe, Cu and Zn were recovered from the leaching solutions by precipitation. The pure metals can be further obtained from precipitates using existing upgrading processes. As a side result, the anion−exchange reactions and metal ions substitution reactions that occurred during the precipitation of Cu and Zn were studied. Furthermore, based on the literature, a satisfactory explanation can be offered for the observation made by XRD.

ASSOCIATED CONTENT Supporting Information Development of chip packaging types; chips remained on the PCBs after 2 hours of leaching with spent TSS; metal content in solution; leaching of waste PCBs with spent TSS.

AUTHOR INFORMATION Corresponding Author


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*

E-mail address: [email protected]; Tel.: +86-10-62794143, Fax: +86-10-

62772048. Notes The authors declare no competing financial interest.

ACKNOWLEDGMENTS This work was supported by the National Key Technology R&D Program (2014BAC03B04). We also thank Dr. Xianlai Zeng and Prof Ab Stevels for their valuable advices. We are also very grateful to Dr. Abhishek Kumar Awasthi for reviewing the grammar of the manuscript.

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For Table of Contents Use Only

A green process of metal recycling: Co-processing waste printed circuit boards and spent tin stripping solution Congren Yang1, Jinhui Li1, *, Quanyin Tan1, Lili Liu2, Qingyin Dong2 1

State Key Joint Laboratory of Environment Simulation and Pollution Control, School of

Environment, Tsinghua University, Beijing 100084, China 2

Basel Convention Regional Centre for Asia and the Pacific, Beijing 100084, China

* Corresponding Author E-mail address: [email protected] (J. Li) Tel.: +86-10-62794143, Fax: +86-10-62772048

TOC

H2SO4 NaOH

pH 1.5

Sn

PCB

SO 42>1.4 Pb 2+

Pb

Spent TSS

Fe

pH 3

Waste PCBs

ECs

Cu Zn

pH 6

pH 8

Synopsis A green co-processing technology is proposed to recover valuable metals from waste printed circuit boards and spent tin stripping solution at room temperature.


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Green Process of Metal Recycling: Coprocessing Waste Printed

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