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Article Hydrometallurgical Process for Recovery from Epoxy-Coated Solid Electrolyte Tantalum Capacitors

Wei-Sheng Chen 1, Hsing-Jung Ho 1,2,* and Kuan-Yan Lin 1,*

1 Department of Resources Engineering, National Cheng Kung University, No.1 University Road, Tainan City 701, Taiwan; [email protected] 2 Department of Environmental Studies for Advanced Society, Graduate School of Environmental Studies, Tohoku University, Aoba-6-6 Aramaki, Aoba-ku, Sendai 980-8577, Miyagi, Japan * Correspondence: [email protected] (H.-J.H.); [email protected] (K.-Y.L.)

 Received: 29 March 2019; Accepted: 10 April 2019; Published: 14 April 2019 

Abstract: Tantalum is a critical that is widely used in electronic products. The demand for tantalum is increasing, but the supply is limited. As tantalum waste products have increased in Taiwan in recent years, the treatment of spent tantalum capacitors has become necessary and important. The recycling of tantalum from tantalum capacitors will not only decrease pollution from waste, but will also conserve tantalum resources. The tantalum content in epoxy-coated solid electrolyte tantalum capacitors (EcSETCs) is over 40 wt.%. Here, we designed a recycling process that includes pre-treatment, , and extraction to recover tantalum. In the pre-treatment process, epoxy resin and wires were removed. During hydrometallurgical process, pressure leaching by hydrofluoric acid was used to leach tantalum and from solid electrolyte tantalum capacitors (SETCs). During our testing of this proposed process, the acid concentration, reaction time, temperature, and solid–liquid ratio were examined for leaching. After the leaching process, Alamine 336 was used to extract tantalum from the leaching solution. The pH value, extractant concentration, extraction time, and aqueous–organic ratio were investigated. Then, tantalum was stripped using HNO3, and the HNO3 concentration, stripping time, and organic–aqueous ratio were analyzed in detail. Under optimal conditions, the recovery efficiency of tantalum reached over 98%, and a final product of tantalum pentoxide with 99.9% purity was obtained after chemical precipitation and .

Keywords: tantalum capacitors; recycling; hydrometallurgy; pressure leaching; solvent extraction

1. Introduction Tantalum (Ta) is an important metal found in the Earth’s crust, having a high (3017 ◦C), high corrosion resistance to acids at normal temperatures, and high conductivity of heat and electricity. Due to those unique properties, tantalum has become indispensable in current industries. Most tantalum production involves the manufacturing of capacitors (34%), which have higher performance than other types of capacitors. Tantalum is also used as specialty chemicals (11%), sputtering targets (13%), super alloys (23%), products (10%), and carbide products (9%) [1,2]. The Tantalum– International Study Center (TIC) indicates that the global production of tantalum was 2800 tons in 2014, and at that time, it was sourced from 1400 tons of tantalum concentrate, 370 tons of , and 990 tons of scrap, respectively [3]. Due to recent developments, the recycling of tantalum has gradually begun to play an important role in the supplementation of tantalum reserves. Yen et al. noted that Taiwan imported 340 tons of tantalum products in 2013

Materials 2019, 12, 1220; doi:10.3390/ma12081220 www.mdpi.com/journal/materials Materials 2019, 12, 1220 2 of 14 and that the major use of tantalum in Taiwan was in capacitors, accounting for 61.6% of the total use at the time [4]. In 2016, the Customs Administration report indicated that Taiwan imported 354 tons of tantalum product. About 80% of the total imported tantalum products remained in Taiwan. The major use of tantalum, again, was in capacitors, accounting for 67%. However, it must be noted that these tantalum products will all eventually become tantalum waste. Because tantalum waste has dramatically increased in recent years, the tantalum recycling process must be studied. Additionally, tantalum always contains a considerable amount of niobium, and a lot of energy is required to remove niobium during the tantalum refining process [5]. The problem is that niobium and tantalum have similar chemical characteristics, which makes refinement difficult. However, the tantalum used in capacitors is free of niobium. Hence, studying the recovery of tantalum from capacitors is not only beneficial for waste management, but it also avoids the disadvantages of the typical tantalum recovery process. Epoxy-coated solid electrolyte tantalum capacitors (EcSETCs) consist of epoxy resin, wires, and electrodes. Epoxy resin is made of halogenated compounds to which silica is added to enhance its thermal durability. The wires are made of and . Electrodes, including cathodes, anodes, and dielectrics, are made of tantalum powder and other elements. Anodes and dielectrics are made of tantalum, and cathodes are made of , paste, and graphite, which are daubed on the surface of the electrode as a cathode layer. EcSETCs are sealed with epoxy resin, which reduces the efficiency of tantalum recovery; hence, the epoxy resin must be removed before any hydrometallurgical processes may be conducted. Many methods can decompose epoxy resin, such as combustion [6], solubilization [7], and super critical water treatment [8]. Those methods have been proposed in previous studies, and steam gasification [9] has been investigated to prevent the generation of harmful halogen compounds in combustion. In the hydrometallurgical process, tantalum capacitors should be dissolved by leaching before purification. In the traditional leaching process, leaching agents, such as, H2SO4 [10,11], hydrofluoric acid (HF) [12,13], or H2SO4 mixed with HF [14], at normal temperatures have been investigated but have been found to achieve only low leaching efficiency. In addition, Wang et al. [15] applied alkali leaching with roast-water, and Rodriguez et al. [16] applied pressure leaching using hydrofluoric acid. Both methods reached high leaching efficiency. Hence, in this study, the leaching process will be improved by combining the advantages of both methods. With respect to solvent extraction, several extractants, such as MIBK [17–25], TBP [19,26,27], CHO [19,22], OCL [28], and Alamine 336 [27–32] have been tested. Previous studies have indicated that Alamine 336 is a promising extractant for purifying tantalum in low HF concentrations [32]. Hussaini et al. indicated that tantalum could be stripped using ammonium or nitric acid in Alamine 336 [31]. Due to the increasing demand for tantalum, the recovery of tantalum from EcSETCs provides new opportunities and alternatives for conservation and waste disposal in Taiwan. As discussed previously, many tantalum capacitors have remained in Taiwan, whereas there are few tantalum reserves. Hence, in this study, we designed a promising recycling process to treat tantalum waste and recover tantalum to obtain a secondary source of this important resource. In our proposed method, epoxy resin and wire are removed by pre-treatment before conducting a hydrometallurgical process. After pre-treatment, EcSETCs without epoxy resin, called solid electrolyte tantalum capacitors (SETCs), are obtained. Then, the SETCs are leached by acid via a leaching process. Subsequently, separation and purification processes are conducted. Here, parameters such as pH value, extractant concentration, extraction time and aqueous-organic ratio were investigated to optimize the extraction of tantalum during this process. Furthermore, during the stripping process, parameters such as concentration, reaction time, and organic–aqueous ratio were studied to achieve high stripping efficiency. After chemical precipitation and calcination, tantalum pentoxide is obtained as the final product of our proposed method. The optimized conditions in this study can be used to further develop the tantalum recovery process in future work. Materials 2019, 12, 1220 3 of 14

2. Materials and Methods

2.1. Materials The tantalum capacitors used in the experiment were EcSETCs from a Taiwanese manufacturer, with 220 µF/10 V. The EcSETCs were divided into three main components: wires, epoxy resin, and electrodes. The relative mass ratio of each component of the EcSETCs is shown in Table1, and the concentration of elements in the EcSETCs is shown in Table2. As shown in Table1, the EcSETCs had 54.8 wt.% electrodes, 38.2 wt.% epoxy resin, and 7 wt.% wires. Table2 shows the results of a chemical analysis of EcSETCs by inductively coupled plasma optical emission spectrometry (ICP-OES, Varian, Vista-MPX). The dominant component of the EcSETCs was tantalum, which was over 40 wt.%, and the other components included 6–8 wt.% manganese, 6–7% , 3% iron, and 4% nickel.

Table 1. Mass ratio of each component in the epoxy-coated solid electrolyte tantalum capacitors (EcSETC) (wt.%).

Tantalum Electrode Epoxy Resin Wire 54.8% 38.2% 7%

Table 2. Concentration of elements in the EcSETCs.

Element Ta Si Mn Fe Ni Wt.% 40–42% 6–7% 6–8% 3% 4%

2.2. Methods The epoxy resin and wires had to be removed from the EcSETCs to raise the leaching efficiency. Hence, pre-treatment was necessary. Pyrolysis with gas was shown to be capable of removing the epoxy resin from the EcSETCs, and it also prevented the generation of halogen gas. Thermogravimetric analysis (TGA, Perkin Elmer, Pyris Diamond TG/DTA) was used in the pre-treatment process to determine the temperature for pyrolysis. Then, grinding and washing with water was conducted to flush out the epoxy resin, which mainly contained silicon . After the washing process, crushing with a (Yuanjiou, YJ-T5520S) was conducted to enhance the efficiency of the leaching process due to the resulting increased contact surface. Finally, magnetic separation was used to remove iron and nickel from the wires. After the pre-treatment process, the resulting SETCs without epoxy resin were analyzed by X-ray diffraction (XRD; DX-2700, Dangdong, Liaoning, China), scanning electron microscopy (SEM; S-3000N, Hitachi, Tokyo, Japan), and energy dispersive X-ray spectrometry (EDS; XFlash6110, Bruker, Billerica, MA, USA). During the leaching process, the SETCs were dissolved using inorganic acid. Different inorganic acids, such as HF (UniRrgion Bio-tech, 49%), nitrate acid (HNO Sigma-Aldrich HNO , 65%), 3 3 ≥ hydrochloric acid (HCl Sigma-Aldrich 37%), and (H2SO4 Sigma-Aldrich 98%), were compared to select the best leaching agent. Parameters such as leaching with or without vapor pressure, acid concentration, temperature, reaction time, and solid–liquid ratio were investigated. The concentration was set from 1% to 49% v/v with vapor pressure at 23 atm and normal vapor pressure. The effect of concentration from 1% to 15% v/v, temperature from 100 to 220 ◦C, reaction time from 0.5 to 3 h, and a liquid–solid ratio from 5 to 100 mL/g were tested to enhance the leaching efficiency. After the leaching process, solvent extraction was used to separate tantalum and other elements to increase the tantalum purity. TBP, Alamine 336, and MIBK diluted in kerosene were used as extractants for tantalum, and their extraction efficiency was analyzed. Then, parameters such as pH values from 1 to 5 in aqueous media, extractant concentration from 0.005 to 0.1 mol/L, aqueous–organic ratio from 0.5 to 9, and reaction time from 0.5 to 5 min were investigated in detail. Next, the organic phase with tantalum was stripped using HNO3. Parameters such as concentration from 0.5 to 3 mol/L, Materials 2019, 12, 1220 4 of 14 organic–aqueous ratio from 1 to 7, and stripping time from 0.5 to 5 min were investigated. Finally, after chemical precipitation and a calcination process, tantalum pentoxide was obtained as the final product.Materials Figure 2019, 121, xshows FOR PEER a flow REVIEW chart of the proposed tantalum recovery process from EcSETCs.4 of 14

FigureFigure 1. 1.The The proposed proposed tantalum tantalum recovery recovery processprocess fromfrom EcSETCs.EcSETCs. ICP-OES: ICP-OES: inductively inductively coupled coupled plasmaplasma optical optical emission emission spectrometry; spectrometry; SEM: SEM: scanningscanning electronelectron microscopy; EDS: EDS: energy energy dispersive dispersive X-rayX-ray spectrometry; spectrometry; TG TG/DTG:/DTG: Thermogravimetric Thermogravimetric analysis. analysis.

3. Results and Discussion

3.1. Pretreatment The EcSETCs were composed of epoxy resin, wires, and electrodes. Epoxy resin and wires connected with electrodes would decrease the leaching efficiency. Th epoxy resin was made of halogenated compounds and silicon oxide; on the other hand, the wire was made of iron and nickel.

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3. Results and Discussion

3.1. Pretreatment The EcSETCs were composed of epoxy resin, wires, and electrodes. Epoxy resin and wires Materials 2019, 12, x FOR PEER REVIEW 5 of 14 connected with electrodes would decrease the leaching efficiency. Th epoxy resin was made of Thesehalogenated elements compounds would have and also silicon affected oxide; the on purity the other of the hand, final the product. wire was Thus, made we of designed iron and nickel.a pre- Thesetreatment elements process would to enhance have the also leaching affected efficiency the purity and of the the purity final of product. the final Thus,product. we designed a pre-treatmentThe TG/DTG process patterns to enhance are shown the leaching in Figure effi ciency2, which and demonstrate the purity of that the finalthe epoxy product. resin began decomposingThe TG/DTG at 350 patterns °C. By the are time shown the in temperatur Figure2, whiche reached demonstrate 600 °C, the that EcSETCs the epoxy had resinlost most began of theirdecomposing mass because at 350 the◦C. epoxy By the resin time had the decomposed temperature completely. reached 600 According◦C, the EcSETCs to results had of the lost TG/DTG most of analysis,their mass 600 because °C for the 10 epoxy min resinwas selected had decomposed as the optimal completely. pyrolysis According condition to results in the of theexperiment. TG/DTG Grindinganalysis, 600and◦ Cwashing for 10 min with was water selected was conducted as the optimal after pyrolysis the pyrolysis condition process. in the Table experiment. 3 shows that Grinding silica couldand washing be removed with water by the was pyroly conductedsis process. after the pyrolysisThen, crushing process. was Table conducted3 shows that before silica magnetic could be separationremoved by to the ensure pyrolysis that process. tantalum Then, and crushing the othe wasr conducted would before be distributed magnetic separationuniformly. to Table ensure 4 showsthat tantalum that iron and and the othernickel metals from would the wires be distributed could be uniformly. removed Tablevia 4magn showsetic that separation. iron and nickelAfter subsequentfrom the wires pre-treatment could be removed processes, via silicon, magnetic nickel, separation. and iron After were subsequent removed. Figure pre-treatment 3 shows processes,SEM and EDSsilicon, analysis nickel, patterns and iron of were the SETCs. removed. As Figurethe result3 showss show, SEM tantalum and EDS and analysis manganese patterns were of distributed the SETCs. Asuniformly the results after show, the pre-treatment. tantalum and In manganese other words, were it distributedwould have uniformly been hard after to recover the pre-treatment. tantalum from In theother SETCs words, directly. it would However, have been hardhere, tothe recover hydromet tantalumallurgical from theprocess SETCs successfully directly. However, separated here, and the recoveredhydrometallurgical tantalum processand manganese. successfully separated and recovered tantalum and manganese.

Figure 2. TGTG/DTG/DTG patterns of the EcSETCs with N22..

Table 3. Removed percentage of silicon in the sample after pyrolysis, grinding, and washing.

TableTable3 3 SiSi Ta Ta BeforeBefore 400 400 mg mg/kg/kg 4500 4500 mg/kg mg /kg AfterAfter 0.2 0.2 mg mg/kg/kg 4500 4500 mg/kg mg /kg RemoveRemove percentage percentage 99.9% 99.9% 0% 0%

Table 4. Remove percentage of nickel and iron after magnetic separation.

Table4 Table 4 NiNi Fe Fe Ta Ta Mag. 99.5% 99.5% 0% Mag. 99.5% 99.5% 0% Non-Mag. 0.5% 0.5% 100% Non-Mag. 0.5% 0.5% 100% Removed percentage 99.5% 99.5% 0% Removed percentage 99.5% 99.5% 0%

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Figure 3. SEM-EDS analysis patterns of the solid electrolyte tantalum capacitors (SETCs) after pre-treatment. FigureFigure 3. SEM-EDS 3. SEM-EDS analysis analysis patterns patterns of of the the solid solid electrelectrolyte tantalumtantalum capacitors capacitors (SETCs) (SETCs) after after pre- pre- 3.2. Leachingtreatment.treatment.

3.2. Leaching3.2.The Leaching SETC powders were leached using four kinds of inorganic acids after pre-treatment. Table5 shows that HF had the highest leaching efficiency with respect to tantalum of the four inorganic The TheSETC SETC powders powders were were leached leached using using four four kindskinds of inorganicinorganic acids acids after after pre-treatment. pre-treatment. Table Table acids, because tantalum could be treated with hydrofluoric acid to produce water-soluble 5 shows5 shows that that HF HFhad had the the highest highest leaching leaching efficiency efficiency withwith respectrespect to to tantalum tantalum of ofthe the four four inorganic inorganic fluorides. Therefore, HF was chosen as the leaching agent to leach tantalum from the SETC powders in acids,acids, because because tantalum tantalum could could be be treated treated with with hydr hydrofluoricofluoric acidacid to to produce produce water-soluble water-soluble hydrogen hydrogen the experiment.. Therefore, HF was chosen as the leaching agent to leach tantalum from the SETC powders fluorides. Therefore, HF was chosen as the leaching agent to leach tantalum from the SETC powders in the experiment. in the experiment. Table 5. Leaching efficiency with respect to tantalum of the different inorganic acids. Table 5. Leaching efficiency with respect to tantalum of the different inorganic acids. Table 5. Leaching efficiencyTable5 with respect to Leachingtantalum of E ffitheciency different of Ta inorganic (%) acids. Table 5 Leaching Efficiency of Ta (%) HydrofluoricHydrofluoricTable acid 5 acid (HF) (HF) Leaching Efficiency 17.57% 17.57% of Ta (%) HydrochloricHydrofluoricHydrochloric acid acid acid (HCl) (HF) (HCl) 17.57%0.13% 0.13% Sulfuric acid (H SO ) 0.01% HydrochloricSulfuric acid acid2 (H (HCl)42SO4) 0.01% 0.13% Nitric acid (HNO3) 0.15% SulfuricNitric acid acid (H (HNO2SO43) 0.15%0.01% (Concentration: 1% v/v; reaction time: 3 h; temperature: 25 C; liquid–solid ratio: 100 mL per gram.) (Concentration: 1% vNitric/v; reaction acid time:(HNO 3 3h;) temperature: 25◦ ℃0.15%; liquid–solid ratio: 100 mL per gram.) (Concentration: 1% v/v; reaction time: 3 h; temperature: 25 ℃; liquid–solid ratio: 100 mL per gram.) TheThe effect effect of vapor of vapor pressure pressure on on the the leaching leaching ofof TaTa waswas investigated. Figure Figure 4 4demonstrates demonstrates that that 5% v/v of HF was enough to leach Ta with 23 atm vapor pressure; on the other hand, leaching with 5%The v/v effect of HF of wasvapor enough pressure to leach on theTa with leaching 23 atm of vapor Ta was pressure; investigated. on the otherFigure hand, 4 demonstrates leaching with that normalnormal pressure pressure required required 49% 49%v/v ofv/v HF. of HF. According According tothe to the results, results, pressure pressure leaching leaching was was used used to to leach 5% v/v of HF was enough to leach Ta with 23 atm vapor pressure; on the other hand, leaching with Ta inleach this study.Ta in Duringthis study. the During pressure the leaching pressure test, leaching parameters test, parameters such as hydrofluoric such as concentration, acid normal pressure required 49% v/v of HF. According to the results, pressure leaching was used to temperature,concentration, reaction temperature, time, and reaction solid-to-liquid time, and ratio solid-to-liquid were investigated. ratio were investigated. leach Ta in this study. During the pressure leaching test, parameters such as hydrofluoric acid concentration, temperature, reaction time, and solid-to-liquid ratio were investigated.

Figure 4. Effect of vapor pressure and normal pressure on tantalum leaching efficiency. Liquid–solid ratio: 100 mL per gram; reaction time: 3 h.

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MaterialsFigure2019Figure ,4.12 Effect4., 1220 Effect of of vapor vapor pressure pressure and and no normalrmal pressurepressure on tantalumtantalum leac leachinghing efficiency. efficiency. Liquid–solid Liquid–solid 7 of 14 ratio:ratio: 100 100 mL mL per per gram; gram; reaction reaction time: time: 3 3 h. h.

3.2.1.3.2.1.3.2.1. EEffectff Effectect of of HydrofluoricHydrofluoric Hydrofluoric AcidAcid Acid ConcentrationConcentration Concentration TheTheThe eeffectff ecteffect ofof hydrofluoricof hydrofluoric hydrofluoric acid acid acid concentration concentration concentration on on the the leaching leachingleaching of Ta ofof andTa Ta and Mnand Mn wasMn was investigatedwas investigated investigated from 1%fromfrom to 1% 15% 1% to vto15%/v 15%at v/ 220v/ atv at ◦220C. 220 The°C. °C. Theresults The results results are are shownare shownshown in Figurein Figure5. 5. The5. TheThe leaching leaching leaching eefficiency ffiefficiencyciency of of of Ta TaTa and andand Mn MnMn increasedincreasedincreased asas as thethe the HFHF HF concentrationconcentration concentration increasedincreased increased when whwhenen HFHF concentrationconcentration varied variedvaried from fromfrom 1% 1%1% to toto 5% 5%5% v/ vv/./v vThis.. ThisThis − + maymaymay havehave have beenbeen been becausebecause because anan an increasedincreased increased amount amount amount of ofof F FF−− andand H+ ionsionsions werewere were available available available to to react react with with the the SETCs SETCs toto formformto form solublesoluble soluble complexes,complexes, complexes, whichwhich which couldcould could be bebe dissolved dissolveddissolved inin water.water.water. The TheThe highest highesthighest leaching leachingleaching efficiency eefficiencyfficiency was waswas achievedachievedachieved whenwhen when thethe the HFHF HF concentrationconcentration concentration reachedreached reached 5% 5%5% v vv///vv.. Hence,Hence, 5% 5%5% vv///vvv of ofof HF HFHF was waswas chosen chosenchosen as asas the the optimal optimal concentrationconcentrationconcentration inin in thethe the experiment.experiment. experiment.

Figure 5. Effect of hydrofluoric acid concentration on the leaching efficiency. Temperature: 220 °C (23 FigureFigureatm). 5. Liquid–solid EffectEffect of of hydrofluoric hydrofluoric ratio: 100 mL acid acid per concentration concentrationgram. Reaction on on time: the the leaching 3 leachingh. efficiency. efficiency. Temperature: Temperature: 220 220 °C (23◦C (23atm). atm). Liquid–solid Liquid–solid ratio: ratio: 100 100mL mLper pergram. gram. Reaction Reaction time: time: 3 h. 3 h. 3.2.2. Effect of Temperature and Reaction Time 3.2.2. Effect of Temperature and Reaction Time 3.2.2. Effect of Temperature and Reaction Time The effect of the reaction time on the leaching efficiency of Ta and Mn at different temperatures The effect of the reaction time on the leaching efficiency of Ta and Mn at different temperatures correspondingThe effect of tothe different reaction vapor time pressureson the leaching was studie efficiencyd. Figure of Ta 6a and shows Mn the at differenteffect of thetemperatures reaction corresponding to different vapor pressures was studied. Figure6a shows the e ffect of the reaction time correspondingtime at different to differenttemperatures vapor on pressuresthe Ta leaching was studieefficiency.d. Figure Figure 6a 6b showsshows the effecteffect of of the the reaction reaction at different temperatures on the Ta leaching efficiency. Figure6b shows the e ffect of the reaction time at timetime at differentat different temperatures temperatures on on the the Ta Mn leaching leaching efficiency. efficiency. Figure The results6b shows show the that effect the of Ta the and reaction Mn different temperatures on the Mn leaching efficiency. The results show that the Ta and Mn leaching timeleaching at different efficiencies temperatures at 220 °C onwere the faster Mn leachingthan those efficiency. at 180 °C. The Ta and results Mn requiredshow that 3 hthe to Tareach and the Mn efficiencies at 220 C were faster than those at 180 C. Ta and Mn required 3 h to reach the highest leachinghighest efficiencies leaching◦ efficiency at 220 °C in were 180 °C faster but onlythan 2those h in◦ 220at 180 °C. °C. Nevertheless, Ta and Mn both required temperatures 3 h to reach could the leaching efficiency in 180 C but only 2 h in 220 C. Nevertheless, both temperatures could achieve highestachieve leaching almost efficiency the same◦ inleaching 180 °C efficiency.but only 2Th ◦h us,in 220we °C.could Nevertheless, choose the appropriateboth temperatures parameters could almostaccording the same to different leaching process efficiency. requirements Thus, we and could different choose demands the appropriate of energy consumption, parameters according process achieve almost the same leaching efficiency. Thus, we could choose the appropriate parameters to dicost,fferent and leaching process requirementsrate. In this study, and we di ffchoseerent 180 demands °C for 3 ofh as energy the optimal consumption, temperature process and reaction cost, and according to different process requirements and different demands of energy consumption, process leachingtime. rate. In this study, we chose 180 C for 3 h as the optimal temperature and reaction time. cost, and leaching rate. In this study, we chose◦ 180 °C for 3 h as the optimal temperature and reaction time.

Figure 6. (a)Effect of the reaction time at different temperatures on Ta leaching efficiency. (b)Effect of the reaction time at different temperatures on Mn leaching efficiency. Concentration: 5 % v/v. Liquid–solid ratio: 100 mL per gram.

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Figure 6. (a) Effect of the reaction time at different temperatures on Ta leaching efficiency. (b) Effect of the reaction time at different temperatures on Mn leaching efficiency. Concentration: 5 % v/v. Materials 2019, 12, 1220 8 of 14 Liquid–solid ratio: 100 mL per gram.

3.2.3.3.2.3. EEffectffect ofof Liquid-to-SolidLiquid-to-Solid RatioRatio TheThe eeffectffect ofof thethe liquid-to-solidliquid-to-solid ratioratio fromfrom 55 toto 100100 mLmL/g/g waswas studied.studied. FigureFigure7 7 shows shows that that the the leachingleaching eefficiencyfficiency waswas increasedincreased whenwhen thethe liquid–solidliquid–solid ratioratio increasedincreased fromfrom 55 toto 3030 mLmL/g./g. ThisThis waswas because,because, whenwhen the the liquid–solid liquid–solid ratio ratio was was low, low, there there was was insu ffiinsufficientcient acid toacid react to inreact this process.in this process. Hence, theHence, optimal the optimal liquid–solid liquid–solid ratio was rati identifiedo was identified to be 30 to mL be/g. 30 mL/g.

FigureFigure 7.7. EEffectffect ofof thethe liquid–solidliquid–solid ratioratio onon thethe leachingleaching eefficiency.fficiency. ReactionReaction time:time: 33 h.h. Concentration:Concentration: 5%5% vv//vv.. Temperature:Temperature: 180180 ◦°C.C.

BasedBased onon thethe results,results, the the optimal optimal conditions conditions for for enhancing enhancing leaching leaching effi efficiencyciency were were determined determined to beto 5%be 5%v/v vof/vHF of HF in a in liquid–solid a liquid–solid ratio ratio of 30 ofmL 30 /mL/gg at 180 at 180◦Cfor °C 3for h. 3 With h. With the optimalthe optimal conditions, conditions, the leachingthe leaching efficiency efficiency of Ta of and Ta Mnand couldMn could reach reach 99%. 99%. Moreover, Moreover, compared compared with with previous previous research research [16], with[16], thewith same the concentrationsame concentration of tantalum-rich of tantalum-ric powder,h powder, the HF consumptionthe HF consumption in this study in this was study reduced was byreduced almost by 70%, almost showing 70%, increasedshowing increased efficiency. efficiency Moreover,. Moreover, the leaching the e ffileachingciency achievedefficiency in achieved this study in wasthis almoststudy was 19% almost higher 19% than higher those achieved than those in previousachieved studies.in previous The reasonstudies. for The this reason was attributed for this was to theattributed fact that to most the offact the that impurities most of in the the impuri samplesties were in removedthe samples during were the removed pre-treatment during process. the pre- treatment process. 3.3. Solvent Extraction and Stripping 3.3. SolventAfter the Extraction leaching and process, Stripping tantalum and manganese were dissolved in hydrofluoric acid. In order to separate tantalum and manganese, a solvent extraction process was conducted. In the solvent After the leaching process, tantalum and manganese were dissolved in hydrofluoric acid. In extraction experiment, the concentrations of Ta and Mn were diluted to 1000 mg/L and 150 mg/L, order to separate tantalum and manganese, a solvent extraction process was conducted. In the solvent respectively. Here, hydrofluoric acid was used as the leaching agent and H SO was used to adjust the extraction experiment, the concentrations of Ta and Mn were diluted to2 10004 mg/L and 150 mg/L, pH value. A comparison of the extractants, the effect of the pH value, the extractant concentration, the respectively. Here, hydrofluoric acid was used as the leaching agent and H2SO4 was used to adjust aqueous-to-organic ratio, and the extraction time were investigated. the pH value. A comparison of the extractants, the effect of the pH value, the extractant concentration, 3.3.1.the aqueous-to-organic Comparison of Extractants ratio, and and the Eextractionffect of pH time Value were investigated.

3.3.1.The Comparison effect of di offferent Extractants extractants and onEffect the of extraction pH Value effi ciency were tested at pH values from 1 to 5. Figure8 shows that Alamine 336 had the highest extraction e fficiency with respect to tantalum of the The effect of different extractants on the extraction efficiency were tested at pH values from 1 to three extractants. This was because Alamine 336 in HF-H2SO4 solution could extract tantalum at a low 5. Figure 8 shows that Alamine 336 had the highest extraction efficiency with respect to tantalum of concentration. Figure9 shows that the extraction e fficiency of Ta decreased as the pH value increased. the three extractants. This was because Alamine 336 in HF-H2SO4 solution could extract tantalum at On the other hand, Mn was not extracted at any pH value. Hence, Alamine 336 and pH 1 were chosen a low concentration. Figure 9 shows that the extraction efficiency of Ta decreased as the pH value as the extractant and optimal pH value in the solvent extraction experiment. increased. On the other hand, Mn was not extracted at any pH value. Hence, Alamine 336 and pH 1 were chosen as the extractant and optimal pH value in the solvent extraction experiment.

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Figure 8. Effect of different extractants on the tantalum extraction efficiency. Concentration: 0.01 mol/L. Extraction time: 5 min. A/O = 1. Figure 8. Effect of different extractants on the tantalum extraction efficiency. Concentration: 0.01 FigureFigure 8.8.E Effectffect of of di ffdifferenterent extractants extractants on theon tantalumthe tantal extractionum extraction efficiency. efficien Concentration:cy. Concentration: 0.01 mol 0.01/L. mol/L. Extraction time: 5 min. A/O = 1. Extractionmol/L. Extraction time: 5 min.time: A5 /min.O = 1.A/O = 1.

Figure 9. Effect of different pH values on the extraction efficiency. Concentration: 0.01 mol/L. Extraction time: 5 min. A/O = 1. FigureFigure 9.9.E ffEffectect of diofff differenterent pH valuespH values on the on extraction the extr eactionfficiency. efficiency. Concentration: Concentration: 0.01 mol/ L.0.01 Extraction mol/L. Figure 9. Effect of different pH values on the extraction efficiency. Concentration: 0.01 mol/L. time:Extraction 5 min. time: A/O 5= min.1. A/O = 1. 3.3.2. ExtractionEffect of Extractant time: 5 min. Concentration, A/O = 1. Aqueous-to-Organic Ratio, and Extraction Time 3.3.2.3.3.2.Figure EEffectffect 10 ofof ExtractantshowsExtractant that Concentration,Concentration, the extraction efficiencyAqueous-to-OrganicAqueous-to-Organic of Ta increased Ratio,Ratio, as andand the ExtractionExtractionAlamine 336 TimeTime concentration 3.3.2. Effect of Extractant Concentration, Aqueous-to-Organic Ratio, and Extraction Time increased from 0.005 to 0.025 mg/L. The extraction efficiency of Ta was highest at a concentration of FigureFigure 1010 showsshows thatthat thethe extractionextraction eefficiencyfficiency ofof TaTa increasedincreased asas thethe AlamineAlamine 336336 concentrationconcentration 0.025 Figuremg/L, which10 shows was that chosen the extraction to be the optimalefficiency extractant of Ta increased concentration as the Alamine in the experiment. 336 concentration Figure increasedincreased fromfrom 0.0050.005 toto 0.0250.025 mgmg/L./L. TheThe extractionextraction eefffificiencyciency ofof TaTa waswas highesthighest atat aa concentrationconcentration ofof 11increased shows that from the 0.005 Ta extraction to 0.025 mg/L. efficiency The extractiondecreased efasficiency the aqueous-to-organic of Ta was highest ratio at a increased concentration from of2 0.0250.025 mgmg/L,/L, whichwhich was was chosen chosen to to be be the the optimal optimal extractant extractant concentration concentration in thein the experiment. experiment. Figure Figure 11 to0.025 9. Hence, mg/L, wewhich chose was an chosen aqueous-to-organic to be the optimal ratio extractantof 2 as the concentrationoptimal aqueous-to-organic in the experiment. ratio Figurein the shows11 shows that that the the Ta extractionTa extraction effi efficiencyciency decreased decrease asd theas the aqueous-to-organic aqueous-to-organic ratio ratio increased increased from from 2 to 2 experiment.11 shows that Figure the Ta 12 extraction shows the efficiency effect of extraction decreased time as the from aqueous-to-organic 0.5 to 5 min. The ratio results increased show that from the 2 9.to Hence,9. Hence, we we chose chose an an aqueous-to-organic aqueous-to-organic ratio ratio of of 2 2 as as the the optimal optimal aqueous-to-organic aqueous-to-organic ratioratio inin thethe extractionto 9. Hence, efficiency we chose of an Ta aqueous-to-organic increased from 0.5 ratio to1 ofmin 2 as and the the optimal reaction aqueous-to-organic was balanced after ratio 1 inmin. the experiment.experiment. Figure 1212 showsshows thethe eeffectffect ofof extractionextraction timetime fromfrom 0.50.5 toto 55 min.min. TheThe resultsresults showshow thatthat thethe Hence,experiment. in order Figure to avoid 12 shows too muchthe effect energy of extraction consumption, time from we chose 0.5 to 15 minmin. as The the results optimal show extraction that the extractionextraction eefficiencyfficiency of of Ta Ta increased increased from from 0.5 to10.5 minto1 min and theand reaction the reaction was balanced was balanced after 1 min.after Hence,1 min. time.extraction efficiency of Ta increased from 0.5 to1 min and the reaction was balanced after 1 min. inHence, order in to order avoid to too avoid much too energy much consumption, energy consumption, we chose we 1 min chose as the 1 min optimal as the extraction optimal time.extraction Hence, in order to avoid too much energy consumption, we chose 1 min as the optimal extraction time. time.

Figure 10. Effect of different concentrations of Alamine 336 on the extraction efficiency. pH = 1. Extraction time: 5 min. A/O = 1.

Materials 2019, 12, x FOR PEER REVIEW 10 of 14 Materials 2019, 12, x FOR PEER REVIEW 10 of 14 Figure 10. Effect of different concentrations of Alamine 336 on the extraction efficiency. pH = 1. Figure 10. Effect of different concentrations of Alamine 336 on the extraction efficiency. pH = 1. MaterialsExtraction2019, 12, 1220time: 5 min. A/O = 1. 10 of 14 Extraction time: 5 min. A/O = 1.

Figure 11. Effect of different aqueous–organic ratios on the extraction efficiency. pH 1. Extraction Figure 11.11. EEffectffect ofof di differentfferent aqueous–organic aqueous–organic ratios ratios on on the the extraction extraction efficiency. efficien pHcy. 1.pH Extraction 1. Extraction time: time: 5 min. Concentration: 0.025 mol/L. time:5 min. 5 Concentration:min. Concentration: 0.025 0.025 mol/L. mol/L.

FigureFigure 12.12. EEffectffectof ofextraction extractiontime timeon on extraction extraction e effifficiency.ciency. pH pH 1. 1. AA/O/O = = 2.2. Concentration:Concentration: 0.025 M. Figure 12. Effect of extraction time on extraction efficiency. pH 1. A/O = 2. Concentration: 0.025 M. Thus, the extraction efficiency of Ta was 99.5% and the extraction efficiency of Mn was 0.1% in the Thus, the extraction efficiency of Ta was 99.5% and the extraction efficiency of Mn was 0.1% in optimalThus, conditions the extraction of 0.025 efficiency mol/L Alamine of Ta was 336 at99.5% pH 1, an ford 1the min, extraction and in an efficiency aqueous-to-organic of Mn was ratio 0.1% of in 2. the optimal conditions of 0.025 mol/L Alamine 336 at pH 1, for 1 min, and in an aqueous-to-organic theIn solvent optimal extraction conditions process, of 0.025 the mol/L separation Alamine effect 336 between at pH 1, Ta for and 1 min, Mn was and demonstrated in an aqueous-to-organic successfully. ratio of 2. In solvent extraction process, the separation effect between Ta and Mn was demonstrated ratio of 2. In solvent extraction process, the separation effect between Ta and Mn was demonstrated successfully. successfully.3.3.3. Effect of Stripping Efficiency on Agent Concentration, Organic-to-Aqueous Ratio and Stripping Time 3.3.3. Effect of Stripping Efficiency on Agent Concentration, Organic-to-Aqueous Ratio and 3.3.3. Effect of Stripping Efficiency on Agent Concentration, Organic-to-Aqueous Ratio and StrippingAfter Time extraction, tantalum was stripped back to the aqueous phase from the organic phase viaStripping exposure Time to HNO3. Next, the concentration, stripping time, and organic–aqueous ratio were After extraction, tantalum was stripped back to the aqueous phase from the organic phase via investigated.After extraction, Figure 13tantalum shows thatwas thestripped effect back of di tofferent the aqueous concentrations phase from of HNO the 3organicon the phase stripping via eexposurefficiency. Asto theHNO results3. Next, show, the the strippingconcentration, efficiency stripping of tantalum time, increased and organic–aqueous as the HNO concentration ratio were exposure to HNO3. Next, the concentration, stripping time, and organic–aqueous3 ratio were increasedinvestigated. from Figure 0.5 to 13 2 molshows/L, andthat nothe increase effect of in different efficiency concentrations was observed of when HNO the3 on concentration the stripping investigated. Figure 13 shows that the effect of different concentrations of HNO3 on the stripping wasefficiency. higher thanAs the 2 mol results/L. Hence, show, the the optimal stripping HNO efficiencyconcentration of tantalum was 2 mol increased/L. Figure as14 showsthe HNO that3 efficiency. As the results show, the stripping 3efficiency of tantalum increased as the HNO3 concentration increased from 0.5 to 2 mol/L, and no increase in efficiency was observed when the concentrationthe stripping eincreasedfficiency offrom tantalum 0.5 to 2 decreased mol/L, and as no the increase organic–aqueous in efficiency ratio was increased observed from when 1to the 7. Asconcentration the results show,was higher an organic–aqueous than 2 mol/L. Hence, ratio ofthe 1 optimal was optimal HNO in3 concentration the stripping was process. 2 mol/L. Figure Figure 15 concentration was higher than 2 mol/L. Hence, the optimal HNO3 concentration was 2 mol/L. Figure 14 shows that the stripping efficiency of tantalum decreased as the organic–aqueous ratio increased 14shows shows that that the the stripping stripping effi efficiencyciency of tantalumof tantalum increased decreased asthe as the stripping organi timec–aqueous increased ratio from increased 0.5 to from 1 to 7. As the results show, an organic–aqueous ratio of 1 was optimal in the stripping process. from5 min. 1 Afterto 7. As 5 min, the results the reaction show, was an organic–aqueou balanced. Finally,s ratio a stripping of 1 was effi optimalciency ofintantalum the stripping of 99.4% process. was Figure 15 shows that the stripping efficiency of tantalum increased as the stripping time increased Figuredemonstrated 15 shows under that thethe optimalstripping stripping efficiency parameters: of tantalum 2 molincreased/L HNO as3 thewith stripping an O/A ratiotime ofincreased 1 and a from 0.5 to 5 min. After 5 min, the reaction was balanced. Finally, a stripping efficiency of tantalum fromstripping 0.5 to time 5 min. of 5 After min. 5 min, the reaction was balanced. Finally, a stripping efficiency of tantalum of 99.4% was demonstrated under the optimal stripping parameters: 2 mol/L HNO3 with an O/A ratio of 99.4% was demonstrated under the optimal stripping parameters: 2 mol/L HNO3 with an O/A ratio of 1 and a stripping time of 5 min. of 1 and a stripping time of 5 min.

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Figure 13. Effect of different concentrations of HNO3 on the stripping efficiency. Stripping time: 10 Figure 13. Effect of different concentrations of HNO3 on the stripping efficiency. Stripping time: 10 Figure 13. Effect of different concentrations of HNO on the stripping efficiency. Stripping time: 10 min. min.Figure O/A 13. = Effect1. of different concentrations of HNO3 3 on the stripping efficiency. Stripping time: 10 min. O/A = 1. Omin./A = O/A1. = 1.

Figure 14. Effect of different organic–aqueous ratios on stripping efficiency. Stripping time: 10 min. FigureFigure 14. 14.E Effectffect of of di differentfferent organic–aqueous organic–aqueous ratios ratios on on stripping stripping e ffiefficiency.ciency. Stripping Stripping time: time: 10 10 min. min. Concentration:Figure 14. Effect 2 mol/L of different HNO3 .organic–aqueous ratios on stripping efficiency. Stripping time: 10 min. Concentration:Concentration: 2 2 mol mol/L/L HNO HNO3.3. Concentration: 2 mol/L HNO3.

Figure 15. Effect of different stripping times on stripping efficiency. O/A = 1. Concentration: 2 mol/L FigureFigure 15. 15.E Effectffect of of di differentfferent stripping stripping times times on on stripping stripping effi efficiency.ciency. O /O/AA = =1. 1. Concentration: Concentration: 2 2 mol mol/L/L HNOFigure3.. 15. Effect of different stripping times on stripping efficiency. O/A = 1. Concentration: 2 mol/L HNO3 3. HNO3. 3.4. Chemical Precipitation and Calcination 3.4.3.4. Chemical Chemical Precipit Precipitationation and and Calcination Calcination 3.4. ChemicalAfter the Precipit solventation extraction, and Calcination stripping process, and selective precipitation, tantalum had already AfterAfter the the solvent solvent extraction, extraction, stri strippingpping process, process, and and selective selective prec precipitation,ipitation, tantalum tantalum had had already already been separatedAfter the solvent from several extraction, impurities. stripping Then, process, during and the selective chemical prec precipitationipitation, tantalum process, had tantalum already beenbeen separated separated from from several several impurities. impurities. Then, Then, duri duringng the the chemical chemical precipitation precipitation process, process, tantalum tantalum wasbeen precipitated separated from by ammonium several impurities. hydroxide Then, at pH duri valueng ofthe 8 tochemical obtain tantalumprecipitation hydroxide. process, Thus, tantalum as a waswas precipitated precipitated by by ammonium ammonium hydroxide hydroxide at at pH pH valu value eof of 8 8to to obtain obtain tantalum tantalum hydroxide. hydroxide. Thus, Thus, as as a a resultwas precipitated ofof thethe calcination calcination by ammonium process, process, we hydroxide we obtained obtained at Ta pH2 OTa5 valuas2O the5 eas of final the 8 to product.final obtain product. tantalum According According hydroxide. to the finalto Thus,the product final as a result of the calcination process, we obtained Ta2O5 as the final product. According to the final productanalysis,result of analysis, thethe crystalline calcination the crystalline phase process, is demonstrated phase we obtained is demo in nstratedTa Figure2O5 as16 in ,the andFigure final the purity16,product. and of the TaAccording2 purityO5 was of proven toTa the2O5 to wasfinal be product analysis, the crystalline phase is demonstrated in Figure 16, and the purity of Ta2O5 was overproduct 99.9% analysis, by ICP-OES. the crystalline phase is demonstrated in Figure 16, and the purity of Ta2O5 was provenproven to to be be over over 99.9% 99.9% by by ICP-OES. ICP-OES. proven to be over 99.9% by ICP-OES.

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FigureFigure 16.16. The XRD patterns for Ta22O55.. 4. Conclusions 4. Conclusions As the demand for tantalum gradually increases, the issue of how to mitigate the exploitation As the demand for tantalum gradually increases, the issue of how to mitigate the exploitation of of this natural resources has become very important. Hence, in this study, we designed and tested this natural resources has become very important. Hence, in this study, we designed and tested a a recycling process for EcSETCs. The raw materials in this study were EcSETCs, which contained recycling process for EcSETCs. The raw materials in this study were EcSETCs, which contained over over 40 wt.% of tantalum. According to this composition, the EcSETCs had great potential for being 40 wt.% of tantalum. According to this composition, the EcSETCs had great potential for being recycled, which could provide a plentiful secondary tantalum source. During this recycling process, the recycled, which could provide a plentiful secondary tantalum source. During this recycling process, EcSETCs were subjected to a pre-treatment, pressure leaching, solvent extraction, stripping, chemical the EcSETCs were subjected to a pre-treatment, pressure leaching, solvent extraction, stripping, precipitation, and calcination. The composition of the EcSETCs was 54.8 wt.% electrode, 38.2 wt.% chemical precipitation, and calcination. The composition of the EcSETCs was 54.8 wt.% electrode, epoxy resin, and 7 wt.% wire. In the pre-treatment process, the epoxy resin was removed by pyrolysis 38.2 wt.% epoxy resin, and 7 wt.% wire. In the pre-treatment process, the epoxy resin was removed at 600 C for 10 min, and then, the resulting SETCs were subjected to grinding and washing by water. by pyrolysis◦ at 600 °C for 10 min, and then, the resulting SETCs were subjected to grinding and The wires, which were made of iron and nickel, were also removed by magnetic separation. The washing by water. The wires, which were made of iron and nickel, were also removed by magnetic optimal pressure leaching conditions were determined to be 5% v/v of HF in a liquid-to-solid ratio of separation. The optimal pressure leaching conditions were determined to be 5% v/v of HF in a liquid- 30 mL/g at 180 C for 3 h. With the optimal conditions, the leaching efficiency of Ta and Mn could to-solid ratio of ◦30 mL/g at 180 °C for 3 h. With the optimal conditions, the leaching efficiency of Ta achieve 99%. In the solvent extraction process, Ta and Mn were separated under 0.025 mol/L Alamine and Mn could achieve 99%. In the solvent extraction process, Ta and Mn were separated under 0.025 336 at pH 1 for 1 min with an aqueous-to-organic ratio of 2. The extraction efficiency of Ta was mol/L Alamine 336 at pH 1 for 1 min with an aqueous-to-organic ratio of 2. The extraction efficiency 99.5%; however, the extraction efficiency of Mn was only 0.1%. In the stripping process, under the of Ta was 99.5%; however, the extraction efficiency of Mn was only 0.1%. In the stripping process, optimal stripping parameters—2 mol/L HNO3 at an organic–aqueous ratio of 1 with a stripping time under the optimal stripping parameters—2 mol/L HNO3 at an organic–aqueous ratio of 1 with a of 5 min—the stripping efficiency of Ta was 99.4%. Then, after chemical precipitation and calcination, stripping time of 5 min—the stripping efficiency of Ta was 99.4%. Then, after chemical precipitation the final product of Ta2O5 was obtained; moreover, the purity of Ta2O5 was over 99.9%, and the and calcination, the final product of Ta2O5 was obtained; moreover, the purity of Ta2O5 was over recovery efficiency of tantalum reached over 98% as a result of our proposed recovery process. Thus, 99.9%, and the recovery efficiency of tantalum reached over 98% as a result of our proposed recovery our proposed recycling process was proven to be successful and effective, and the detailed reaction process. Thus, our proposed recycling process was proven to be successful and effective, and the mechanism and optimization of this process will be discussed in future work. detailed reaction mechanism and optimization of this process will be discussed in future work. Author Contributions: W.-S.C. and K.-Y.L. conceived of and designed the experiments; K.-Y.L. performed and Authoranalyzed Contributions: the experiments; W.-S.C. H.-J.H. and conducted K.-Y.L. conceived and edited of and the designed research structure;the experiments; and K.-Y.L. K.-Y.L. and performed H.-J.H. wrote and analyzedthe paper. the experiments; H.-J.H. conducted and edited the research structure; and K.-Y.L. and H.-J.H. wrote theFunding: paper. This research was funded by NCKU Research & Development Foundation (106S281). Acknowledgments:Funding: This researchWe was are pleasedfunded toby acknowledge NCKU Research the support& Development of the Laboratory Foundation of Resources (106S281). Circulation (LRC) at National Cheng Kung University. Acknowledgments: We are pleased to acknowledge the support of the Laboratory of Resources Circulation (LRC)Conflicts at National of Interest: ChengThe Kung authors University. declare no conflicts of interest.

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