Cleaner Extraction of Lead from Complex Lead-Containing Wastes by Reductive Sulfur-Fixing Smelting with Low SO2 Emission

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Article Cleaner Extraction of Lead from Complex Lead-Containing Wastes by Reductive Sulfur-Fixing Smelting with Low SO2 Emission

Yun Li 1,2, Shenghai Yang 1, Wenrong Lin 1,3, Pekka Taskinen 2, Jing He 1, Yuejun Wang 4, Junjie Shi 2, Yongming Chen 1,* , Chaobo Tang 1,* and Ari Jokilaakso 1,2,*

1 School of Metallurgy and Environment, Central South University, Changsha 410083, China; yun.li@aalto.fi (Y.L.); [email protected] (S.Y.); [email protected] (W.L.); [email protected] (J.H.) 2 Department of Chemical and Metallurgical Engineering, School of Chemical Engineering, Aalto University, 02150 Espoo, Finland; pekka.taskinen@aalto.fi (P.T.); junjie.shi@aalto.fi (J.S.) 3 Zhuhai Coslight Battery Co., Ltd, Zhuhai 519180, China 4 Department of Ecology and Resources Engineering, Hetao College, Bayannur 015000, China; [email protected] * Correspondence: [email protected] (Y.C.); [email protected] (C.T.); ari.jokilaakso@aalto.fi (A.J.); Tel.: +86-731-88830470 (Y.C. & C.T.); +358-50-3138885 (A.J.)

Received: 31 December 2018; Accepted: 7 February 2019; Published: 17 February 2019

Abstract: A novel and cleaner process for lead and silver recycling from multiple lead-containing wastes, e.g., lead ash, lead sludge, lead slag, and ferric sludge, by reductive sulfur-fixing smelting was proposed. In this process, coke and iron-containing wastes were employed as reductive agent and sulfur-fixing agent, respectively. A Na2CO3-Na2SO4 mixture was added as flux. The feasibility of this process was detected from thermodynamic and experimental perspectives. The influence of Fe/SiO2 and CaO/SiO2, composition of the molten salt, coke addition, smelting temperature, and smelting time on direct Pb recovery and sulfur-fixation efficiency were investigated. The optimal process

conditions were determined as follows: WCoke = 15% WPb wastes, WNa2CO3 /WNa2SO4 = 0.7/0.3, Fe/SiO2 ◦ = 1.10, CaO/SiO2 = 0.30, smelting temperature 1200 C, and smelting time 2 h, where W represents weight. Under these optimum conditions, 92.4% Pb and 98.8% Ag were directly recovered in crude lead bullion in one step treatment, and total 98.6% sulfur was ﬁxed. The generation and emissions of SO2 can be avoided. The main phases in ferrous matte obtained were FeS, NaFeS2, Fe2Zn3S5, and a little entrained Pb. The slag was a FeO-SiO2-CaO-Na2O quaternary melt.

Keywords: SO2-free sulfur ﬁxing smelting; lead recycling; wastes co-treatment and recycling; lead-iron-containing wastes; molten salt; PbO-FeO-SiO2-CaO-Na2O phase diagram

1. Introduction Today, large amounts of lead-containing wastes are produced in non-ferrous metallurgical industry [1–3], especially in Pb and Zn metallurgy fields, such as lead ash and lead slag generated in Pb/Zn smelting, lead anode slime, and lead sludge produced in the electrolytic refining. Other engineering fields [4–6] are also sources of Pb-bearing substances, including lead scrap and lead paste separated from spent lead-acid batteries [7]. In many countries, these lead-bearing wastes are classified as hazardous waste due to the high toxicity of lead [8]. They are greatly detrimental to environment and human health if left untreated and abandoned directly to the environment [9,10]. However, these wastes generally contain considerable amounts of precious metals, such as gold (Au) and silver (Ag). It also enormously challenges the circular economy if the valuable metals are not recycled. Various lead wastes are potential resources for extracting lead and precious metals

Minerals 2019, 9, 119; doi:10.3390/min9020119 www.mdpi.com/journal/minerals Minerals 2019, 9, 119 2 of 15 considering the fact that the grades of lead concentrates are decreasing [11] and primary lead ores are gradually exhausting. Many studies about the treatment of Pb-containing wastes using pyrometallurgical and hydrometallurgical processes were reported in recent years [12,13]. The immobilization of lead slags to prepare construction materials and inorganic polymers were also investigated [14,15]. Hydrometallurgical methods also can be applied for various Pb-bearing wastes treatment. They often consist of a two-step process of leaching and separation in acid [16], alkaline [17], and other lixiviants [18,19]. Many hydrometallurgical processes involve shortcomings of a long process chain, large amounts of reagents consumption, and waste water generation. Most hydrometallurgical disposal processes of lead wastes are still far from the industrial production scale. Pyrometallurgy is the predominant methodology around the world for primary lead production. Reverberatory smelting and bath smelting are the two main technologies [20]. Generally, sulfur content in the feed required in the reverberatory smelting techniques is lower than 2%. As a result, SO2 emission is inevitable with Pb-bearing wastes because sulfur contents of lead wastes typically are higher than 2%. Thus, a high energy consumption and heavy pollution are involved when using the reverberatory smelting technology. At present, many lead and zinc smelters apply bath smelting processes to treat Pb wastes along with the primary raw materials [21], such as spent lead-acid battery pastes, lead slags, and other ferrous wastes. However, compared to the primary sulfide concentrate, the sulfur content in wastes is too low to provide enough heat for the smelting when considering the endothermic decomposition reactions of sulfate [22]. Therefore, the use of waste materials often leads to an unstable smelting operation. Moreover, some new technologies [23–25] such as vacuum separation have been reported to treat lead-bearing residues and lead paste. They need a desulfurization pre-treatment step, and the operation cost is relatively high. Alkaline smelting was first employed by researchers of the former Soviet Union to extract lead from lead sulfide using molten NaOH [26]. Subsequently, more molten salt systems were further developed and introduced to extract lead [27], antimony [28], and bismuth [29] due to the high solvent property, thermal stability, and electrical property of molten salt [30]. In the molten salt smelting, the traditional FeO-SiO2-CaO slag will be replaced by a molten salt. In view of environment protection and comprehensive utilization of secondary resources, this investigation proposed a promising new process, referred to as reducing sulfur-fixing smelting, to efficiently recover Pb, Ag and recycle iron and sulfur values from lead-bearing wastes in an environmentally friendly manner. Na2CO3-Na2SO4 binary salt was used as reaction medium. The novelty of this process is treatment of various iron-lead-bearing wastes, SO2-free sulfur fixation, a much shorter flowsheet, absence of harmful by-product, and a wide adaptability for different secondary materials. This work investigated the experimental feasibility of this new process in detail. The effects of slag property, feed mixture of different raw materials, and smelting temperature and time on lead recovery and sulfur-fixation efficiency were determined. Furthermore, a series of bench-pilot experiments were conducted. The distributions of Pb, Fe, and S were also evaluated.

2. Experimental

2.1. Materials The lead-bearing wastes used in this study come from the processes of Chinese metallurgical industry. The feed mixture contained lead ash, lead sludge, lead slag, and ferric sludge. In order to have a steady assay of the feed, 23.9 g lead ash, 48.0 g lead sludge, 5.0 g lead slag, and 23.1 g ferric sludge were manually and carefully mixed to obtain a mixture containing 30% of Pb and 8% of S. Coke used as the reducing agent in this research had ﬁxed carbon of 84.3%. The chemical compositions of the materials are given in Table1. Figure1 illustrates the X-ray diffraction analysis results (XRD, D/max 2550PC, Rigaku Co. Ltd, Tokyo, Japan). Other reagents, including Na2CO3 and Na2SO4, ﬂux Fe2O3, SiO2, and CaO, were analytical purity and supplied by Aladdin Industrial Corporation, China. Minerals 2019, 9, 119 3 of 15

Minerals 2019, 9, x FOR PEER REVIEW 3 of 16 Table 1. Chemical compositions of the materials and their mixture (wt.%). Table 1. Chemical compositions of the materials and their mixture (wt.%). Materials Pb S Fe Zn Ag * SiO2 CaO Materials Pb S Fe Zn Ag * SiO2 CaO Lead ash 49.08 6.74 3.78 2.22 45 0.89 0.39 Lead ash 49.08 6.74 3.78 2.22 45 0.89 0.39 Lead sludge 30.51 12.19 3.77 6.72 65 5.93 8.61 Lead sludgeLead slag30.51 18.5612.19 0.723.77 8.896.72 0.42 3565 44.135.93 0.79 8.61 Lead slagIron sludge18.56 11.610.72 2.178.89 40.07 0.42 0.36 73035 6.8444.13 7.83 0.79 Iron sludgeMixture 11.61 30.002.17 8.0040.07 12.41 0.36 4.2 236730 7.196.84 6.07 7.83 Mixture 30.00 8.00 12.41 4.2 236 7.19 6.07 Industrial Analysis/% Chemical Composiztion of the Ash/% Coke Industrial Analysis/% Chemical Composiztion of the Ash/% Cfix*Vd*Ad* Fetotal SiO2 CaO Al2O3 Coke Cfix* Vd* Ad* Fetotal SiO2 CaO Al2O3 81.27 3.30 15.43 25.23 41.23 6.60 25.24 81.27 3.30 15.43 25.23 41.23 6.60 25.24 * g/t; Cfix: fixed carbon; Vd: dry basis volatiles; Ad: dry basis ash. * g/t; Cfix: fixed carbon; Vd: dry basis volatiles; Ad: dry basis ash.

1600 8 -PbS 9-Pb3O4 .-Fe2(SO4)3 -ZnSO4 1400 8,9 -PbO·PbSO4 8 1200 Lead ash  1000 8 . 9  9.  8 8 9 .   8 8 8 800   

-PbSO -CaSO ·2H O 1 1 1 4 4 2

Intensity[A.U.] 600 2,6 3-Pb Zn ((SO )(SiO )(Si O )) 2 3 1 4 2 4 4 2 7 4-Zn O(SO ) 5-PbO 1 3 4 2 2 400 5 1 6-Fe6(OH)12CO3 1 1 2 Lead sludge 200 1 1 1 1 1 1 5 1 3 6 2 1 4 3 2 4 6 5 5 0 10 20 30 40 50 60 70 80 2θ/[○]

6 500 1-Fe2O3 3-Pb8Ca(Si2O7)3 -PbSiO3

7-PbO(Fe2O3)6 8-FeSiO3 9-Fe3O4  7 400 8 6 9 7 7 3 7 6 7 1 Lead slag 3 6 8 6 6 6 7 7 69 7 7 9 300 1

-Fe2O3 2-2PbO·PbSO4 3-Pb8Ca(Si2O7)3 200 4-PbO·2Fe2O3 5-ZnSO3·2H2O Intensity [A.U.] Intensity 2 3,4 Iron sludge 100 1 5 1 3 3 2 4 1 3 1 1 3 1 1 3 1 1 2 4 1 1 2 3 2 4 1 0

10 20 30 40 ○ 50 60 70 80 2θ/[ ]

Figure 1. XRD patterns of the feed materials. The XRD spectra shows that the main lead-bearing phases in the feed materials were sulfate The XRD spectra shows that the main lead-bearing phases in the feed materials were sulfate (PbSO4), sulﬁde (PbS), oxide (Pb3O4 and PbO2), and silicate (PbSiO3). At the same time, zinc sulfate (PbSO4), sulfide (PbS), oxide (Pb3O4 and PbO2), and silicate (PbSiO3). At the same time, zinc sulfate (ZnSO4), zinc oxysulfate (Zn3O(SO4)2), iron sulfate (Fe2(SO4)3), calcium sulfate (CaSO4), a complex (ZnSO4), zinc oxysulfate (Zn3O(SO4)2), iron sulfate (Fe2(SO4)3), calcium sulfate (CaSO4), a complex lead-zinc mineral queitite (Pb4Zn2((SO4)(SiO4)(Si2O7)), iron silicate (FeSiO3), and ferric oxide (Fe2O3) lead-zinc mineral queitite (Pb4Zn2((SO4)(SiO4)(Si2O7)), iron silicate (FeSiO3), and ferric oxide (Fe2O3) were also detected in the materials. The feed material is thus a polymetallic complex material. were also detected in the materials. The feed material is thus a polymetallic complex material.

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Minerals2.2. Methods 2019, 9 , x 119 FOR PEER REVIEW 44 of of 16 15

2.2. MethodsFigure 2 shows a detailed flowchart of this novel process. As an experimental procedure, 100 g 2.2.lead Methods-bearing waste mixture was mixed with coke powder, Na2CO3-Na2SO4 mixture, and other fluxes (CaO,Figure SiO2 and2 shows Fe2O a3 ifdetailed necessary). flowchart The mixture of this novelwas placed process. in anAs alumina an experimental crucible. procedure,The crucible 100 was g Figure2 shows a detailed flowchart of this novel process. As an experimental procedure, 100 g leadpushed-bearing into the waste constant mixture temperature was mixed zone with of acoke tube powder, furnace atNa the2CO desired3-Na2SO temperature.4 mixture, andThe otherschematic fluxes of lead-bearing waste mixture was mixed with coke powder, Na2CO3-Na2SO4 mixture, and other fluxes (CaO,the tube SiO furnace2 and Fe is2O illustrated3 if necessary). in Figure The mixture 3. In the was smelting placed process in an alumina, a series crucible. of reduction The crucible and sulfur was- (CaO, SiO2 and Fe2O3 if necessary). The mixture was placed in an alumina crucible. The crucible was pushedfixation into reactions the constant took place temperature between zone the ofreactants a tube furnace. Metallic at thePb wasdesired extracted temperature. from the The raw schematic materials of pushed into the constant temperature zone of a tube furnace at the desired temperature. The schematic theand tube enriched furnace in crudeis illustrated lead bullion. in Figure At the 3. Insame the time,smelting the sprocessulfur was, a seriesfixed asof ferrousreduction matte, and insteadsulfur- of the tube furnace is illustrated in Figure3. In the smelting process, a series of reduction and fixationof emitting reactions SO2. After took theplace required between smelting the reactants time in. M air,etallic the cruciblePb was extractedwas taken from out fromthe raw the materials hot zone sulfur-fixation reactions took place between the reactants. Metallic Pb was extracted from the raw andand enrichedcooled to in room crude temperature. lead bullion. Next, At the the same crucible time, was the broken sulfur towas carefully fixed as separate ferrous andmatte, we insteadigh the materials and enriched in crude lead bullion. At the same time, the sulfur was fixed as ferrous matte, ofsmelting emitting products SO2. After obtained, the required crude smelting lead, ferrous time in matte air, the, and crucible slag. Thewas productstaken out werefrom the analyzed hot zone by instead of emitting SO2. After the required smelting time in air, the crucible was taken out from the hot andInductively cooled to Coupled room temperature. Plasma-Atomic Next, Emission the crucible Spectrometry was broken (ICP to- AES,carefully Optima separate 3000, and Perkin weigh Elmer, the zone and cooled to room temperature. Next, the crucible was broken to carefully separate and weigh smeltingNorwalk, productsCT, USA) obtained, and X-ray crude diffractometer. lead, ferrous matte, and slag. The products were analyzed by the smelting products obtained, crude lead, ferrous matte, and slag. The products were analyzed by Inductively Coupled Plasma-Atomic Emission Spectrometry (ICP-AES, Optima 3000, Perkin Elmer, Inductively Coupled Plasma-Atomic Emission Spectrometry (ICP-AES, Optima 3000, Perkin Elmer, Norwalk, CT, USA) and X-ray diffractometer. Norwalk, CT, USA) and X-ray diffractometer.

Figure 2. Process flow chart of the reductive sulfur-fixing smelting. Figure 2. Process ﬂow chart of the reductive sulfur-ﬁxing smelting. Figure 2. Process flow chart of the reductive sulfur-fixing smelting.

Figure 3. A schematic of the experimental apparatus. Figure 3. A schematic of the experimental apparatus.

Figure 3. A schematic of the experimental apparatus.

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The direct Pb recovery η and sulfur-ﬁxing rate ξ were calculated using the following equations:

Mass of lead in the crude lead Direct Pb recovery η = (1) Mass of lead in the initial feed materials Mass of sulfur in the matte and slag Sulfur − ﬁxation rate ξ = (2) Mass of sulfur in the initial feed materials

3. Thermodynamics Consideration

The main lead-bearing phases in the feed materials are lead sulfate (PbSO4), lead sulﬁde (PbS), and lead oxide (PbO). In the smelting process, a series of reduction and sulfur-ﬁxation reactions took place between the reactants at high temperature (1000–1300 ◦C), as shows in Table2.

Table 2. Possible reactions taking place in the reductive sulfur-ﬁxing smelting process of lead-bearing complex mixture.

θ ◦ Reactions ∆GT−T [kJ/mol, C] [31] Equation θ 3Fe2O3 + C = 2Fe3O4 + CO(g) ∆GT = −0.235 T + 74.482 (3) θ 3Fe2O3 + CO(g) = Fe3O4 + CO2(g) ∆GT = −0.061 T − 47.13 (4) θ 3/16 PbSO4 + 1/16 Fe3O4 + C = 3/16 Pb + 3/16 FeS + CO(g) ∆GT = −0.177 T + 61.90 (5) θ ∆GT = −0.004 T − 60.39 T ≤ 860 ◦C 3/16 PbSO4 + 1/16 Fe3O4 + CO(g) = 3/16 Pb + 3/16 FeS + CO2(g) θ (6) ∆GT = 0.004 T − 67.22 T ≥ 860 ◦C θ 3/4 PbS + 1/4 Fe3O4 + C = 3/4 Pb + 3/4 FeS + CO(g) ∆GT = −0.192 T + 117.48 (7) θ 3/4 PbS + 1/4 Fe3O4 + CO(g) = 3/4 Pb + 3/4 FeS + CO2(g) ∆GT = −0.017 T − 5.932 (8) θ 1/5 PbSO4 + 1/5 ZnO + C = 1/5 Pb + 1/5 ZnS + CO(g) ∆GT = −0.172 T + 51.4 (9) θ ∆GT = −0.004 T − 353.56 T ≤ 900 ◦C 1/5 PbSO4 + 1/5 ZnO + CO(g) = 1/5 Pb + 1/5 ZnS + CO2(g) θ (10) ∆GT = −0.0096 T − 79.87 T ≥ 900 ◦C θ PbS + ZnO + C = Pb + ZnS + CO(g) ∆GT = −0.177 T + 83.51 (11) θ ∆GT = −0.007 T− 37.229 T ≤ 1100 ◦C PbS + ZnO + CO(g) = Pb + ZnS + CO2(g) θ (12) ∆GT = −0.031 T− 79.884 T ≥ 1100 ◦C θ 1/6 PbSO4 + 1/6 Na2CO3 + C = 1/6 Pb + 1/6 Na2S + 7/6 CO(g) ∆GT = −0.196 T + 92.02 (13) θ 1/5 PbSO4 + 1/5 Na2CO3 + CO(g) = 1/5 Pb + 1/5 Na2S + 6/5 CO2(g) ∆GT = −0.025 T − 37.68 (14) θ 1/2 PbS+ 1/2 Na2CO3 + C = 1/2 Pb + 1/2 Na2S + 3/2 CO(g) ∆GT = −0.244 T + 189.31 (15) θ PbS + Na2CO3 + CO(g) = Pb + Na2S + 2CO2(g) ∆GT = −0.138 T + 131.78 (16) θ 2/7 Na2SO4 + 2/7 FeS + C = 2/7 NaFeS2 + 1/7 Na2O + CO(g) ∆GT = −0.150 T + 107.28 (17) θ 2/7 Na2SO4 + 2/7 FeS + CO(g) = 2/7 NaFeS2 + 1/7 Na2O + CO2(g) ∆GT = 0.025 T − 16.13 (18) θ PbO + C = Pb + CO(g) ∆GT = −0.178 T + 51.425 (19) θ ∆GT = −0.112 T − 67.781 T ≤ 880 ◦C PbO + CO(g) = Pb + CO2(g) θ (20) ∆GT = 0.014 T − 91.033 T ≥ 880 ◦C

θ The Gibbs energy change ∆GT of all above reactions was calculated using HSC Chemistry 9.2.6 θ and its database [31]. Figure4 shows the ∆GT vs.T diagrams of Reactions (3)–(20). It illustrates that lead extraction reactions in the presence of sulfur-ﬁxing agents and reductant are thermodynamically favorable in the temperature range of 900–1300 ◦C. Increase in temperature will promote the positive trends of the reactions. The reductive sulfur-ﬁxing reactions in this smelting process are feasible. Metallic Pb can be extracted from PbSO4, PbS, and PbO, and enriched in crude lead bullion. At the Minerals 2019, 9, x FOR PEER REVIEW 6 of 16

The Gibbs energy change ∆ of all above reactions was calculated using HSC Chemistry 9.2.6 and its database [31]. Figure 4 shows the ∆ vs. diagrams of Reactions (3)–(20). It illustrates that lead extraction reactions in the presence of sulfur-fixing agents and reductant are thermodynamically Mineralsfavorable2019, 9in, 119 the temperature range of 900–1300 °C. Increase in temperature will promote the positive6 of 15 trends of the reactions. The reductive sulfur-fixing reactions in this smelting process are feasible. Metallic Pb can be extracted from PbSO4, PbS, and PbO, and enriched in crude lead bullion. At the same time, the sulfur in the initial raw materials was transferred to FeS, NaFeS2, ZnS, and ultimately same time, the sulfur in the initial raw materials was transferred to FeS, NaFeS2, ZnS, and ultimately ﬁxed as ferrous matte, instead of emitting SO2. fixed as ferrous matte, instead of emitting SO2.

Figure 4. ∆Gθ -T diagram of Reactions (3)–(20). The data were obtained from HSC Chemistry 9.2.6 and T Figure 4. ∆ -T diagramθ of Reactions (3)–(20). The data were obtainedθ from HSC Chemistry 9.2.6 its database [31]. (a) ∆GT-T diagram of Reactions (5)–(12); (b) ∆GT-T diagram of Reactions (3),(4),(19) and its database [31]. (a) ∆ -T diagram of Reactions (5)–(12); (b) ∆ -T diagram of Reactions and (20); (c) ∆Gθ -T diagram of Reactions (13)–(18). T (3),(4),(19) and (20); (c) ∆ -T diagram of Reactions (13)–(18). 4. Results and Discussion 4. Results and Discussion 4.1. Effect of Fe/SiO2 4.1. Effect of Fe/SiO2 The effect of Fe/SiO2 on the direct Pb recovery and the sulfur-fixing rates was investigated The effect of Fe/SiO2 on the direct Pb recovery and the sulfur-fixing rates was investigated in the in the range from Fe/SiO2 = 0.65 to 1.3 (w/w), with a fixed conditions of Wcoke = 10% range from Fe/SiO2 = 0.65 to 1.3 (w/w), with a fixed conditions of Wcoke = 10% Wraw materials, molten salt Wraw materials, molten salt addition WNa2CO3 + Na2SO4 = 18% Wraw materials and molten salt composition raw materials addition WNa2CO3 Na2SO4 = 18% W and molten salt composition WNa2CO3 / WNa2SO4 = 0.3/0.7◦ WNa2CO3 /WNa2SO4 = 0.3/0.7 (where W represents weight), CaO/SiO2 = 0.3, smelting at 1200 C for(where 2 h. TheW represents results areweight), illustrated CaO/SiO in2 Figure= 0.3, smelting5. It can at be1200 observed °C for 2 h. that The the results direct are Pb illustrated recovery in Figure 5. It can be observed that the direct Pb recovery increased gradually with increasing Fe/SiO2 increased gradually with increasing Fe/SiO2 and peaked at Fe/SiO2 = 1.1 where 85.3% Pb was and peaked at Fe/SiO2 = 1.1 where 85.3% Pb was directly recovered. However, a steady declining directly recovered. However, a steady declining trend was observed when the Fe/SiO2 was above 1.1. trend was observed when the Fe/SiO2 was above 1.1. Otherwise, the sulfur-fixing rate increased Otherwise, the sulfur-fixing rate increased slowly from 89.5% to 95.5% as Fe/SiO2 increased from 0.65 slowly from 89.5% to 95.5% as Fe/SiO2 increased from 0.65 to 1.3, and reached a maximum sulfur- to 1.3, and reached a maximum sulfur-fixing rate of 96.7% at Fe/SiO2 = 1.25. This can be explained by the fact that a large fraction of iron oxide in the slag promotes the sulfur-fixing reactions.

Figure6 illustrates a liquidus surface diagram of the PbO-FeO-SiO 2-CaO-Na2O system with ﬁxed ratio of WNa2O/WSiO2 = 0.5 and WCaO/WSiO2 = 1/3. It reveals that the temperature of liquid phase boundaries deceased with the increasing FeO fraction. This indicates that a suitable FeO addition decreases the melting point of the slag; at the same time, the viscosity and ﬂuidity will also be improved. As a result, the reduction and settling environment of lead particles are improved. However, it is also Minerals 2019, 9, x FOR PEER REVIEW 7 of 16

fixing rate of 96.7% at Fe/SiO2 = 1.25. This can be explained by the fact that a large fraction of iron oxide in the slag promotes the sulfur-fixing reactions.

100

Minerals 2019, 9, 119 7 of 15 90 [%]

observed that excessive FeO fraction will also lead to a higher melt density and higher melting point inMinerals slag. 2019 It is, 9, unbeneﬁcial x FOR PEER REVIEW to the settling and elimination of Pb droplets. At the same time, the7 blueof 16 liquidus lines gradually Percentage moved80 to PbO corner with the increasing FeO fraction. This indicates that the solubilityfixing rate of of lead 96.7% increase at Fe/SiO in the2 = slag. 1.25. Therefore, This can be a moderate explained Sulfur-fixing addition by ratethe fact of iron that helps a large to obtain fraction a suitable of iron 75 FeO-SiOoxide in 2the-CaO-Na slag promotes2O slag for the effective sulfur-fixing lead recoveryreactions. Direct and sulfur-ﬁxation.Pb recovery rate

10070 0.6 0.7 0.8 0.9 1.0 1.1 1.2 1.3 1.4 Fe/SiO 95 2

Figure 5. Effect of Fe/SiO90 2 on the Pb recovery and sulfur-fixing efficiency. (Wcoke = 10% Wraw materials [%] raw materials 2 , WNa2CO3 Na2SO4= 18% W , WNa2CO3 / WNa2SO4 = 0.3/0.7, CaO/SiO = 0.3, 1200 °C, 2 h). 85

Figure 6 illustrates a liquidus surface diagram of the PbO-FeO-SiO2-CaO-Na2O system with fixed ratio of W /W = 0.5 and W /W = 1/3. It reveals that the temperature of liquid phase Na2 Percentage Si80O2 CaO SiO2 boundaries deceased with the increasing FeO fraction. This indicates that a suitable FeO addition Sulfur-fixing rate decreases the melting point75 of the slag; at the same Direct time, Pb recovery the viscosity rate and fluidity will also be improved. As a result, the reduction and settling environment of lead particles are improved. However, it is also observed70 that excessive FeO fraction will also lead to a higher melt density and 0.6 0.7 0.8 0.9 1.0 1.1 1.2 1.3 1.4 higher melting point in slag. It is unbeneficial to the settling and elimination of Pb droplets. At the Fe/SiO same time, the blue liquidus lines gradually moved to 2PbO corner with the increasing FeO fraction. This indicates that the solubility of lead increase in the slag. Therefore, a moderate addition of iron Figure 5. Effect of Fe/SiO2 on the Pb recovery and sulfur-ﬁxing efﬁciency. (Wcoke = 10% Wraw materials, helpsFigure to obtain 5. Effect a suitable of Fe/SiO FeO2- SiOon 2 the-CaO Pb-Na recovery2O slag andfor effective sulfur-fixing lead efficiency. recovery and(Wcoke sulfur = ◦ 10%-fixation. Wraw WNa2CO3 + Na2SO4 = 18% Wraw materials, WNa2CO3 /WNa2SO4 = 0.3/0.7, CaO/SiO2 = 0.3, 1200 C, 2 h). materials raw materials 2 , WNa2CO3 Na2SO4= 18% W , WNa2CO3 / WNa2SO4 = 0.3/0.7, CaO/SiO = 0.3, 1200 °C, 2 h).

Figure 6 illustrates a liquidus surface diagram of the PbO-FeO-SiO2-CaO-Na2O system with

fixed ratio of WNa2 /WSiO2= 0.5 and WCaO /WSiO2 = 1/3. It reveals that the temperature of liquid phase boundaries deceased with the increasing FeO fraction. This indicates that a suitable FeO addition decreases the melting point of the slag; at the same time, the viscosity and fluidity will also be improved. As a result, the reduction and settling environment of lead particles are improved. However, it is also observed that excessive FeO fraction will also lead to a higher melt density and higher melting point in slag. It is unbeneficial to the settling and elimination of Pb droplets. At the same time, the blue liquidus lines gradually moved to PbO corner with the increasing FeO fraction. This indicates that the solubility of lead increase in the slag. Therefore, a moderate addition of iron helps to obtain a suitable FeO-SiO2-CaO-Na2O slag for effective lead recovery and sulfur-fixation.

Figure 6. Liquidus contour diagram of the PbO-FeO-SiO2-CaO-Na2O system, the data was taken from Figure 6. Liquidus contour diagram of the PbO-FeO-SiO2-CaO-Na2O system, the data was taken from MTDATA vers. 8.2 [32] MTOX database [33]. MTDATA vers. 8.2 [32] MTOX database [33].

4.2. Effect of CaO/SiO2 Ratio of the Slag

Figure7 shows the effect of CaO/SiO 2 ratio of the silicate slag on the lead extraction and sulfur ﬁxation. It is observed that the direct Pb recovery dropped sharply from 86.4% to 69.4% when CaO/SiO2 was increased from 0.3 to 0.45. This indicates that the increasing CaO/SiO2 was harmful to the settling of metallic lead. However, the change of CaO/SiO2 had almost no effect on sulfur-ﬁxing rate. It remained nearly constant in around 97%. Figure8 illustrates a liquidus contour diagram of the

Figure 6. Liquidus contour diagram of the PbO-FeO-SiO2-CaO-Na2O system, the data was taken from MTDATA vers. 8.2 [32] MTOX database [33].

MineralsMinerals 20192019,, 99,, xx FORFOR PEERPEER REVIEWREVIEW 88 ofof 1616

4.2. Effect of CaO/SiO2 Ratio of the Slag 4.2. Effect of CaO/SiO2 Ratio of the Slag

Figure 7 shows the effect of CaO/SiO2 ratio of the silicate slag on the lead extraction and sulfur Figure 7 shows the effect of CaO/SiO2 ratio of the silicate slag on the lead extraction and sulfur fixation.fixation. It It is is observed observed that that the the direct direct Pb Pb recovery recovery dropped dropped sharply sharply from from 86.4% 86.4% to to 69.4% 69.4% when when CaO/SiO2 was increased from 0.3 to 0.45. This indicates that the increasing CaO/SiO2 was harmful to CaO/SiO2 was increased from 0.3 to 0.45. This indicates that the increasing CaO/SiO2 was harmful to Mineralsthe settling2019, 9of, 119 metallic lead. However, the change of CaO/SiO2 had almost no effect on sulfur-fixing8 of 15 the settling of metallic lead. However, the change of CaO/SiO2 had almost no effect on sulfur-fixing rate.rate. ItIt remainedremained nearlynearly constantconstant inin aroundaround 9977%.%. FigureFigure 88 illustratesillustrates aa liquidusliquidus contourcontour diagramdiagram ofof the CaO-FeO-SiO2-Na2O system in fixed W /W i ratio of 0.5. It also indicates that increasing the CaO-FeO-SiO2-Na2O system in fixed W Na2 /W Si O2 ratio of 0.5. It also indicates that increasing CaO-FeO-SiO2-Na2O system in ﬁxed WNa2O/NaW2SiO2 ratioS O2 of 0.5. It also indicates that increasing addition additionofaddition CaO will ofof CaOCaO lead will towill an leadlead increase toto anan in increaseincrease melting inin point meltingmelting of the pointpoint slag. ofof As thethe a slag.slag. result, AsAs the aa result,result, viscosity thethe and viscosityviscosity ﬂuidity andand of fluidity of slag will also deteriorate. The Pb recovery showed a diminishing trend at higher CaO fluidityslag will of also slag deteriorate. will also deteriorate. The Pb recovery The Pb showed recovery a diminishing showed a trenddiminishing at higher trend CaO at additions. higher CaO additions. additions. 100100

9595 Sulfur-fixing rate 90 Sulfur-fixing rate 90 Direct Pb recovery rate Direct Pb recovery rate 8585

80 80 Percentage [%] Percentage [%] 7575

7070

6565

0.300.30 0.330.33 0.360.36 0.390.39 0.420.42 0.450.45 CaO/SiO CaO/SiO 2 2 Figure 7. 7. EffectEffect of CaO/SiO of CaO/SiO2 ratio on ratiothe direct on Pb the recover directy Pband recoverysulfur-fixing and efficiency sulfur-ﬁxing.(Wcoke = efﬁciency. 10%Wraw Figure 7. Effect of CaO/SiO2 ratio on2 the direct Pb recovery and sulfur-fixing efficiency.(Wcoke = 10%Wraw (materialsW , W= 10%Na COW Na SO = 18%, WWraw materials+, WNa CO= / 18%WNaWSO = 0.3/0.7, ,Fe/SiOW 2 = 1.1,/W 1200 °C, 2= h) 0.3/0.7,. materialscoke, W 2 3 raw materials2 4 = 18%WNaraw2 materialsCO3 , NaW2SO24 3 / W 2 raw4 = 0.3/0.7, materials Fe/SiONa22 =CO 1.1,3 1200Na2 SO°C,4 2 h). Na2CO3 Na2SO◦ 4 Na2CO3 Na2SO4 Fe/SiO2 = 1.1, 1200 C, 2 h).

Figure 8. Liquidus contour diagram of the CaO-FeO-Na O-SiO system with W /W = 0.5. Figure 8. Liquidus contour diagram of the CaO-FeO-Na22O-SiO2 2system with W Na2O/WSiO2 = 0.5. Figure 8. Liquidus contour diagram of the CaO-FeO-Na2O-SiO2 system with W Na2 /W SiO2= 0.5. The data was taken from MTDATA vers. 8.2 MTOX database. Na2 SiO2 TThehe datadata waswas takentaken fromfrom MTDATAMTDATA vers.vers. 8.28.2 MTOXMTOX database.database. 4.3. Effect of W /W Na2CO3 Na2SO4

The relationship between WNa2CO3 /WNa2SO4 and Pb recovery and sulfur-ﬁxing rates are presented in Figure9. It can be seen that the direct Pb recovery increased steadily from 85.3% to 90.8% as the WNa2CO3 /WNa2SO4 increased from 0.3/0.7 to 0.9/0.1. However, the sulfur-fixing rate was Minerals 2019, 9, x FOR PEER REVIEW 9 of 16 Minerals 2019, 9, x FOR PEER REVIEW 9 of 16

4.3. Effect of WNa2CO3/WNa2SO4 4.3. Effect of WNa2CO3/WNa2SO4

The relationship between WNa2CO3 / WNa2SO4 and Pb recovery and sulfur-fixing rates are Minerals 2019The, 9 , 119relationship between WNa CO / WNa SO and Pb recovery and sulfur-fixing rates are 9 of 15 presented in Figure 9. It can be seen that2 the3 direct2 4Pb recovery increased steadily from 85.3% to 90.8% presented in Figure 9. It can be seen that the direct Pb recovery increased steadily from 85.3% to 90.8% as the WNa2CO3 / WNa2SO4 increased from 0.3/0.7 to 0.9/0.1. However, the sulfur-fixing rate was as the WNa CO / WNa SO increased from 0.3/0.7 to 0.9/0.1. However, the sulfur-fixing rate was constant on2 around3 2 95.5%.4 Molten salt can simultaneously decrease the melting point, density, constantconstant on aroundon around 95.5%. 95.5%. Molten Molten salt can simultaneously simultaneously decrease decrease the the melting melting point, point, density, density, viscosity [34,35], and improve the fluidity of the slag [36]. Therefore, Na2CO3-Na2SO4 mixture can viscosityviscosity[34, 35[34,35]], and, and improve improve the the fluidity fluidity of of the the slagslag [[36]36].. Therefore, Therefore, Na Na2CO2CO3-Na3-Na2SO42 SOmix4turemixture can can provide a flexible reaction medium and significantly promote the smelting reactions. Figure 10 provideprovide a flexible a flexible reaction reaction medium medium and significantly and significantly promote promote the smelting the smelting reactions. reactions. Figure Figure 10 illustrates 10 illustrates the phase diagram of the binary Na2CO3-Na2SO4 system. It shows that the melting the phaseillustrates diagram the phase of the diagram binary of Na the2CO binary3-Na 2 NaSO24COsystem.3-Na2SO4 It system. shows It that shows the meltingthat the temperaturemelting temperature of Na2CO3-Na2SO4 mixture decreases with increasing Na2CO3 fraction, and reaches the temperature of Na2CO3-Na2SO4 mixture decreases with increasing Na2CO3 fraction, and reaches the of Naminimum2CO3-Na 2 inSO W4 mixture/W decreases= 0.54/0.45. with Therefore,increasing with Na2 increasingCO3 fraction, W and/ reaches W theratio, minimum the minimum in W Na2CO3/W Na2SO4= 0.54/0.45. Therefore, with increasing W Na2CO3/ W Na2SO4 ratio, the W W Na2CO3 Na2SO4 W Na2WCO3 Na2SO4 in Nasettling2CO3 / environmentNa2SO4 = 0.54/0.45.of lead particles Therefore, was improved. with increasing More metallicNa Pb2CO was3 / settledNa2SO 4andratio, collected the settling in environmentsettling environment of lead particles of lead was particles improved. was improved. More metallic More metallic Pb was settledPb was andsettled collected and collected in crude in lead. crude lead. crude lead. 100 100

96 96 [%]

[%] 92 92

Percentage 88 Percentage

Sulfur-fixing rate 84 Sulfur-fixing rate 84 Direct Pb recovery rate Direct Pb recovery rate

80 80 0.3/0.7 0.5/0.5 0.7/0.3 0.9/0.1 0.3/0.7 0.5/0.5 0.7/0.3 0.9/0.1 WNa CO / WNa SO WNa CO2 /3 WNa SO2 4 2 3 2 4 Figure 9. Effect of Na2CO3 / Na2SO4 ratio on the Pb recovery and sulfur-fixing efficiency (Wcoke FigureFigure 9. 9Effect. Effect of of Na Na2CO2CO33/Na / Na2SO2SO44 ratio on on the the Pb Pb recovery recovery and and sulfur sulfur-fixing-fixing efficiency efficiency (Wcoke (W coke raw materials raw materials 2 2 . =10%W , WNa2CO3 Na2SO4 = 18%W , Fe/SiO = 1.1, CaO/SiO = 0.3, 1200 °C, 2 h) ◦ =10%Wraw materials, W = 18%Wraw materials, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, 1200 °C, 2 h). = 10%Wraw materials, WNaNa22COCO3 +NaNa2SO2SO4 4 = 18%Wraw materials, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, 1200 C, 2 h).

FigureFigure 10. Phase 10. Phase diagram diagram of theof the Na Na2CO2CO3-Na3-Na22SOSO4 system;system; the the data data was was taken taken from from FactSage FactSage 7.2 and 7.2 its and its Figure 10. Phase diagram of the Na2CO3-Na2SO4 system; the data was taken from FactSage 7.2 and its FTsalt database [37]. FTsaltFTsalt database database [37 ].[37].

4.4. Effect of Coke Addition Figure 11 illustrates the effect of coke addition on the smelting yields. When coke dosage was increased from 5% to 25%, Pb extraction was found to increase from 81.3% to 91.5%, and it reached a maximum of 92.4% at 20% coke addition. The sulfur-fixing rate increased from 91.7% to 94.8%, and peakedMinerals 2019 at 97.2%, 9, 119 with 15% coke addition. However, a small decrease of sulfur-fixation was observed10 of 15 when the coke addition was more than 15%. Strong reductive atmosphere was beneficial to the reduction and enrichment of lead. It was also intended to promote iron oxide reduction. More FexOy was reducedreduced toto generate generate metallic metallic Fe Fe and and then then transferred transferred into into crude crude lead lead phase. phase. Afterwards, Afterwards, iron oxide iron oxideacting acting as sulfur-ﬁxation as sulfur-fixation agent correspondingly agent correspondingly decreased. decreased. As a result, As the a result, slag composition the slag composition was turned wasto a directionturned to thata direction was against that was the Pbagainst recovery, the Pb as shownrecovery, in Figureas shown6. Therefore, in Figure a6. slight Therefore, decrease a slight was decreaseobserved was in both observed the Pb in recovery both the and Pb sulfur-ﬁxationrecovery and sulfur beyond-fixation 15% coke beyond addition. 15% coke addition.

100

92 [%]

84 Percentage Sulfur-fixing rate Direct Pb recovery rate 80

76 5 10 15 20 25 Coke dosage / % Figure 11. W Figure 11. Effect ofof cokecoke dosagedosage onon thethe Pb Pb recovery recovery and and sulfur-fixing sulfur-fixing efficiency efficiency ( (WNaNa2CO2CO3 +3 Na Na22SOSO44 = ◦ 18%WPb materials , W /W = 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, 1200 C, 2 h). 18%WPb materials, WNaNa2CO2CO3 /W3 Na2NaSO24SO= 40.7/0.3, Fe/SiO = 1.1,2 CaO/SiO = 0.3,2 1200 °C, 2 h). 4.5. Effect of Smelting Temperature 4.5. Effect of Smelting Temperature The direct Pb recovery and sulfur-fixing rates at different temperatures are shown in Figure 12. The direct Pb recovery and sulfur-fixing rates at different temperatures are shown in Figure 12. Below 1200 ◦C, Pb recovery and sulfur-fixing rates presented a sharp rising trend with increasing Below 1200 °C, Pb recovery and sulfur-fixing rates presented a sharp rising trend with increasing temperatures. The Pb recovery increased from 58.3% to 93.3% and sulfur-fixing rate increased at the temperatures. The Pb recovery increased from 58.3% to 93.3% and sulfur-fixing rate increased at the same time from 75.9% to 96.2%. The phase diagram of the PbO-FeO-SiO2-CaO-Na2O system shown in same time from 75.9% to 96.2%. The phase diagram of the PbO-FeO-SiO2-CaO-Na2O system shown Figure6 also indicates that increasing temperature is beneficial to decrease the PbO content in the slag. in Figure 6 also indicates that increasing temperature is beneficial to decrease the PbO content in the However, direct Pb recovery steadily decreased when the temperature was higher than 1200 ◦C due to slag. However, direct Pb recovery steadily decreased when the temperature was higher than 1200 °C the intensified volatilization of metallic Pb, but sulfur-fixation continued to increase slightly. Therefore, due to the intensified volatilization of metallic Pb, but sulfur-fixation continued to increase slightly. a suitable smelting temperature was critical to achieve a high direct Pb recovery to the molten metal, Therefore, a suitable smelting temperature was critical to achieve a high direct Pb recovery to the and to avoid lead losses in the off-gas fumes. moltenMinerals 2019metal,, 9, x and FOR toPEER avoid REVIEW lead losses in the off-gas fumes. 11 of 16

100

90 [%] 80

Percentage Sulfur-fixing rate

60 Direct Pb recovery rate

50 1050 1100 1150 1200 1250 1300 Temperature [C]

Figure 12. Effect of temperature on the Pb recovery and sulfur-fixing efficiency (Wcoke = 15%WPb materials, Figure 12. Effect of temperature on the Pb recovery and sulfur-fixing efficiency (Wcoke=15%WPb materials, WNa CO + Na SO = 18%WPb materials, WNa CO /WNa SO = 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, 2 h). 2 3 2 4 Pb materials 2 3 2 4 2 2 WNa2CO3 Na2SO4 = 18%W , WNa2CO3 / WNa2SO4= 0.7/0.3, Fe/SiO = 1.1, CaO/SiO = 0.3, 2 h).

4.6. Effect of Smelting Time Figure 13 shows the effect of smelting time on the Pb recovery and sulfur-fixation. It illustrates that extending the smelting time can promote the recovery of Pb and sulfur fixation. 93.3% Pb and 97.5% sulfur were recovered and immobilized within 2 h of reaction. After 3 h, Pb recovery and sulfur-immobilization rates were 92.6% and 98.3%, respectively. This indicates 2 h reaction time was enough for the settling and enrichment of the Pb product in the bullion. The further extension of smelting time would lead to a gradual decline in Pb recovery due to volatilization of metallic Pb. When the smelting time was increased from 2 h to 3 h, the sulfur-fixing rate showed a slight increase from 97.5% to 98.3%. Considering the volatilization of Pb and energy consumption, the optimum smelting time was suggested to be around 2 h.

100

Percnetage [%] Percnetage Sulfur-fixing rate 80 Pb direct recovery rate

75 0.5 1.0 1.5 2.0 2.5 3.0 Smelting time [h]

Figure 13. Effect of smelting time on the Pb recovery and sulfur-fixing efficiency (Wcoke = 15% WPb materials, Pb materials 2 2 WNa2CO3Na2SO4 = 18%W , WNa2CO3 / WNa2SO4= 0.7/0.3, Fe/SiO = 1.1, CaO/SiO = 0.3, 1200 °C).

4.7. Bench-Pilot Experiments Based on the above detailed experiments in the laboratory, the optimum processing conditions coke raw materials raw materials were obtained as: W /W = 15%, WNa2CO3 Na2SO4 = 18% W , WNa2CO3 / WNa2SO4 = 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, smelting temperature 1200 °C, and smelting time 2 h. These conditions were then applied to carry out two bench-pilot experiments with 200g Pb-bearing

Minerals 2019, 9, x FOR PEER REVIEW 11 of 16

100

90 [%] 80

Percentage Sulfur-fixing rate

60 Direct Pb recovery rate

50 1050 1100 1150 1200 1250 1300 Temperature [C]

MineralsFigure2019, 12.9, 119 Effect of temperature on the Pb recovery and sulfur-fixing efficiency (Wcoke=15%WPb materials11, of 15 Pb materials 2 2 WNa2CO3 Na2SO4 = 18%W , WNa2CO3 / WNa2SO4= 0.7/0.3, Fe/SiO = 1.1, CaO/SiO = 0.3, 2 h).

4.6. Effect of Smelting Time Figure 13 shows the effect of smelting time on the Pb recovery and sulfur-fixation.sulfur-fixation. It illustrates that extending the smelting time can promote the recovery of Pb and sulfursulfur fixation.fixation. 93.3% Pb and 97.5% sulfur were recovered and immobilized within 2 h h ofof reaction.reaction. After 3 h, Pb recovery and sulfur-immobilizationsulfur-immobilization rates were 92.6% and 98.3%, respectively. This indicates 2 h reaction time was enough for the settling and enrichment of the PbPb productproduct inin thethe bullion.bullion. The further extension of smelting time would lead to a gradual decline in Pb recovery due to volatilization of of metallic metallic Pb. Pb. When the smelting time was increased from 2 h to 3 h, the sulfur sulfur-fixing-fixing rate showed a slight increase from 97.5% toto 98.3%.98.3%. Considering the volatilization of Pb and energy consumption,consumption, the optimumoptimum smelting time was suggested to be around 22 h.h.

100

Percnetage [%] Percnetage Sulfur-fixing rate 80 Pb direct recovery rate

75 0.5 1.0 1.5 2.0 2.5 3.0 Smelting time [h]

Figure 1 13.3. EffectEffect of smelting of smelting time on time the onPb recovery the Pb recoveryand sulfur and-fixing sulfur-ﬁxing efficiency (W efﬁciencycoke = 15% W (WPbcoke materials=, WNa CO Na SO = 18%WPb materials, WNa CO / WNa SO = 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, 1200 °C). 15% 2WPb3 materials2 4 , WNa2CO3 + Na2SO4 2= 18%3 WPb2 materials4 , WNa2CO3 /WNa2SO4 = 0.7/0.3, Fe/SiO2 = 1.1, ◦ CaO/SiO2 = 0.3, 1200 C). 4.7. Bench-Pilot Experiments 4.7. Bench-Pilot Experiments Based on the above detailed experiments in the laboratory, the optimum processing conditions Based on the abovecoke raw detailed materials experiments in the laboratory,raw materials the optimum processing were obtained as: W /W = 15%, WNa2CO3 Na2SO4 = 18% W , WNa2CO3 / WNa2SO4 = conditions were obtained as: W /W = 15%, W = 18% W , 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3,coke smeltingraw materials temperature 1200Na 2°C,CO3 and+ Na 2smeltingSO4 time raw2 h. materials These W /W = 0.7/0.3, Fe/SiO = 1.1, CaO/SiO = 0.3, smelting temperature 1200 ◦C, conditionsNa2CO3 wereNa2SO 4 then applied to carry2 out two bench-2pilot experiments with 200g Pb-bearing and smelting time 2 h. These conditions were then applied to carry out two bench-pilot experiments with 200g Pb-bearing materials mixture. The chemical compositions of different products obtained, crude Pb, ferrous matte and slag, were presented in Table3. It was found that lead concentrations in the ferrous matte and slag were below 2.4% and 0.7%, respectively. Purity of the crude lead was higher than 96.5%, and its main impurity was metallic Fe.

Table 3. Chemical compositions of smelting products in the bench-pilot experiments (wt.%).

Crude Pb Ferrous Matte Slag No. Pb Fe Ag * Pb Fe S Zn Na Pb Fe S Zn Na CaO SiO2 1 96.06 2.12 821 2.54 48.80 24.07 2.36 7.47 0.61 23.01 4.21 1.51 8.65 8.59 27.30 2 97.02 1.72 845 2.30 50.19 24.51 3.30 6.94 0.72 24.58 3.07 1.83 8.24 8.97 28.71 AVG. 96.54 1.92 833 2.42 49.50 24.29 2.83 6.71 0.67 23.80 3.64 1.67 8.45 8.78 28.01 * g/t. Minerals 2019, 9, x FOR PEER REVIEW 12 of 16

materials mixture. The chemical compositions of different products obtained, crude Pb, ferrous matte and slag, were presented in Table 3. It was found that lead concentrations in the ferrous matte and slag were below 2.4% and 0.7%, respectively. Purity of the crude lead was higher than 96.5%, and its main impurity was metallic Fe.

Table 3. Chemical compositions of smelting products in the bench-pilot experiments (wt.%).

Crude Pb Ferrous Matte Slag No. Pb Fe Ag* Pb Fe S Zn Na Pb Fe S Zn Na CaO SiO2 1 96.06 2.12 821 2.54 48.80 24.07 2.36 7.47 0.61 23.01 4.21 1.51 8.65 8.59 27.30 2 97.02 1.72 845 2.30 50.19 24.51 3.30 6.94 0.72 24.58 3.07 1.83 8.24 8.97 28.71 AVG. 96.54 1.92 833 2.42 49.50 24.29 2.83 6.71 0.67 23.80 3.64 1.67 8.45 8.78 28.01 Minerals 2019, 9, 119 * g/t. 12 of 15

The optical macrograph and XRD results of smelting products are illustrated in Figure 14. It was observedThe optical that macrographthe product was and separated XRD results to ofthree smelting layers: productsslag, ferrous are illustratedmatte, and in crude Figure lead. 14. Slag It was observedfloats on that the the surface product of the was ferr separatedous matte, to and three metallic layers: Pb slag, subsides ferrous at matte,the bottom and crudedue to lead. their Slagdensity floats ondifferences, the surface which of the guarantees ferrous matte, a good and separation metallic Pb and subsides recovery at during the bottom the smelting due to theirprocess. density In differences,addition, whichthe main guarantees components a good in the separation solidified and matte recovery were FeS, during NaFeS the2 smelting, and Fe2Zn process.3S5. It ind In addition,icated thethat main sulfur components was fixed in in the the solidified matte in form matte of were sulfide FeS, matte. NaFeS A2 little, and entrained Fe2Zn3S5 metallic. It indicated Pb phase that was sulfur wasalso fixed detected in the in matte it. The in formslag was of sulfide a FeO- matte.SiO2-CaO A- littleNa2O entrained quaternary metallic and solidified Pb phase as was a vitreous also detected slag in it.system. The slag was a FeO-SiO2-CaO-Na2O quaternary and solidified as a vitreous slag system.

(A)

FigureFigure 14. 14.(A (A) Physical) Physical macrograph macrograph andand (B) XRD patterns patterns of of the the products products in in the the bench bench-pilot-pilot experiments experiments..

TheThe distribution distribution behaviors behaviors of of Pb, Pb, Ag,Ag, Fe,Fe, andand S in the bench bench-pilot-pilot experiments experiments are are summarized summarized inin Figure Figure 15 15.. As As shown, shown, 92.4% 92.4% Pb Pb couldcould bebe directlydirectly enriched in in the the crude crude lead, lead, and and only only 5.5% 5.5% and and 1.8%1.8% of of Pb Pb distributed distributed into into thethe ferrousferrous matte matte and and slag, slag, respectively. respectively. 98.8% 98.8% of Ag of Agwas was also alsoenriched enriched in incru crudede lead. lead. 71.0% 71.0% of of iron iron was was divided divided into into ferrous ferrous matte matte as sulfide as sulfide and the and rest the of rest 28.7% of 28.7% was in was the in theFeO FeO-SiO-SiO2-CaO2-CaO-Na-Na2O 2slag,O slag, respectively. respectively. Total Total sulfur sulfur-fixed-fixed rate was rate 98. was6% 98.6%to condensed to condensed phases, phases,with with91.3% 91.3% in matte in matte,, 7.3% 7.3%in slag in. At slag. the same At the time, same 0.2% time, sulfur 0.2% was sulfur in crude was lead in. crudeOnly 1.2% lead. distributed Only 1.2% distributedto the fume. to the Therefore, fume. Therefore, this novel this reductive novel reductivesulfur-fixing sulfur-fixing smelting process smelting can process significantly can significantly inhibit inhibitMineralsthe generation the2019 generation, 9, x FOR of PEERSO of2 and REVIEW SO largely2 and largely decrease decrease the SO2the emissions. SO2 emissions. 13 of 16

FigureFigure 15.15. TheThe distributiondistribution behaviors behaviors of of Pb, Pb, Ag, Ag, Fe, Fe and, and S inS in crude crude lead, lead, ferrous ferrous matte, matte, slag, slag and, and off-gas off- fumegas fume in the in bench-pilotthe bench-pilot tests. tests.

Recently, this novel Pb recycling process from polymetallic complex materials was industrially adopted [38] in several recycling plants in China, such as Guoda Nonferrous Metals Metallurgy Co., Ltd. [39] and Yongchang Precious Metals Co., Ltd. It has practically proved feasible. The industrial application results indicate that this process can be steadily executed on a continuous basis and offers advantages of low dust generation and high production capacity in currently available smelting furnaces. The operating expenditure was reduced significantly. The off-gas emissions met the relevant Chinese standards. Furthermore, the matte produced can be sold directly for sulfuric acid manufacture and regenerated as sulfur-fixing agent. The slag is harmless and can be used as a raw material for cement production after water-quenching and granulation. Otherwise, in addition to LAB paste recycling, it also can be used for treating antimony- [28,30], bismuth- [40], zinc- [41], and copper-containing [42] wastes and residues. Sb and Bi are enriched and recovered in the bullion. Zinc and copper are primarily distributed to the sulfide matte. Direct recoveries of the metals, i.e. yields without subsequent slag cleaning and matte processing, reach more than 90%, and the sulfur-fixing rate is generally higher than 95%.

5. Conclusion Cleaner extraction of lead from complex polymetallic lead-containing wastes using reductive sulfur-fixing smelting process was fundamentally and experimentally confirmed to be feasible. The

coke raw materials optimum smelting conditions were determined in this study as: W =15%W , WNa2CO3/ 2 2 WNa2SO4= 0.7/0.3, Fe/SiO = 1.1, CaO/SiO = 0.3, smelting temperature 1200 °C and smelting time 2 h. Under these conditions, three smelting products, crude lead, ferrous matte, and slag, were obtained. While 92.4% Pb was directly recovered as crude lead, 98.8% Ag was enriched in the crude lead. Lead concentration in the produced ferrous matte and slag were less than 2.4% and 0.7%, respectively. Purity of the crude lead was higher than 96.5%. Sulfur was fixed in the matte and slag as sulfide. The total sulfur-fixing rate was 98.3%, including 90.3% in the matte and 8.0% in the slag. Gaseous SO2 generation and emissions could be essentially eliminated. This process is a promising high-efficiency technique that is cleaner and offers multiple application prospects in the fields of lead-, antimony-, bismuth-, zinc-, and copper-containing secondary materials, wastes, and residues recycling.

Author Contributions: Y.L. contributed to the works of literature search, figures, data collection, data analysis and writing—review and editing; S.Y. contributed to the conceptualization and supervising works; W.L. and Y.L. performed the experiments together, and W.L. wrote the original draft; J.H. and Y.W. contributed to the design of the work; J.S. contributed to the data analysis. Y.C., C.T., and A.J. played a contributor role of study design, data interpretation, and project administration; A.J. and

Minerals 2019, 9, 119 13 of 15

Recently, this novel Pb recycling process from polymetallic complex materials was industrially adopted [38] in several recycling plants in China, such as Guoda Nonferrous Metals Metallurgy Co., Ltd. [39] and Yongchang Precious Metals Co., Ltd. It has practically proved feasible. The industrial application results indicate that this process can be steadily executed on a continuous basis and offers advantages of low dust generation and high production capacity in currently available smelting furnaces. The operating expenditure was reduced significantly. The off-gas emissions met the relevant Chinese standards. Furthermore, the matte produced can be sold directly for sulfuric acid manufacture and regenerated as sulfur-fixing agent. The slag is harmless and can be used as a raw material for cement production after water-quenching and granulation. Otherwise, in addition to LAB paste recycling, it also can be used for treating antimony- [28,30], bismuth- [40], zinc- [41], and copper-containing [42] wastes and residues. Sb and Bi are enriched and recovered in the bullion. Zinc and copper are primarily distributed to the sulfide matte. Direct recoveries of the metals, i.e., yields without subsequent slag cleaning and matte processing, reach more than 90%, and the sulfur-fixing rate is generally higher than 95%.

5. Conclusions Cleaner extraction of lead from complex polymetallic lead-containing wastes using reductive sulfur-fixing smelting process was fundamentally and experimentally confirmed to be feasible. The optimum smelting conditions were determined in this study as: Wcoke = 15%Wraw materials, ◦ WNa2CO3 /WNa2SO4 = 0.7/0.3, Fe/SiO2 = 1.1, CaO/SiO2 = 0.3, smelting temperature 1200 C and smelting time 2 h. Under these conditions, three smelting products, crude lead, ferrous matte, and slag, were obtained. While 92.4% Pb was directly recovered as crude lead, 98.8% Ag was enriched in the crude lead. Lead concentration in the produced ferrous matte and slag were less than 2.4% and 0.7%, respectively. Purity of the crude lead was higher than 96.5%. Sulfur was fixed in the matte and slag as sulfide. The total sulfur-fixing rate was 98.3%, including 90.3% in the matte and 8.0% in the slag. Gaseous SO2 generation and emissions could be essentially eliminated. This process is a promising high-efficiency technique that is cleaner and offers multiple application prospects in the fields of lead-, antimony-, bismuth-, zinc-, and copper-containing secondary materials, wastes, and residues recycling.

Author Contributions: Y.L. contributed to the works of literature search, figures, data collection, data analysis and writing—review and editing; S.Y. contributed to the conceptualization and supervising works; W.L. and Y.L. performed the experiments together, and W.L. wrote the original draft; J.H. and Y.W. contributed to the design of the work; J.S. contributed to the data analysis. Y.C., C.T., and A.J. played a contributor role of study design, data interpretation, and project administration; A.J. and P.T. contributed the work of writing—review and editing and final approval of the version to be published. Funding: This work was supported by the Guangdong Provincial Applied Science and Technology Specialized Research Project [Grant No. 2016B020242001], the Hunan Provincial Science Fund for Distinguished Young Scholars [Grant No. 2018JJ1044], the National Natural Science Foundation of China [Grant No. 51234009 and 51604105], CMEco [Grant No. 2116781] by Business Finland, the Fund for Less Developed Regions of the National Natural Science Foundation of China (Grant No. 51664013), the Program for Young Talents of Science and Technology in Universities of the Inner Mongolia Autonomous Region (Grant No. NJYT-17-B35), and the Bayannur Science and Technology Project from the Bayannur Bureau for Science and Technology (Grant No. K201509). Conflicts of Interest: The authors declare no conflict of interest.

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