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Article Extraction of Lead and from a Rotary Kiln Oxidizing Cinder

Junhui Xiao 1,2,3,* , Kai Zou 1,3, Wei Ding 1,3, Yang Peng 1,3 and Tao Chen 1,3

1 Sichuan Engineering Lab of Non-Metallic Powder Modification and High-Value Utilization, Southwest University of and , Mianyang 621010, China; [email protected] (K.Z.); [email protected] (W.D.); [email protected] (Y.P.); [email protected] (T.C.) 2 Key Laboratory of Sichuan Province for Comprehensive Utilization of and Titanium Resources, Panzhihua University, Panzhihua 61700, China 3 Key Laboratory of Ministry of Education for Waste Treatment and Resource Recycle, Southwest University of Science and Technology, Mianyang 621010, China * Correspondence: [email protected] or [email protected]; Tel.: +86-139-9019-0544

 Received: 28 February 2020; Accepted: 1 April 2020; Published: 2 April 2020 

Abstract: In this study, sulfuric leaching and gravity shaking-table separation by shaking a table are used to extract lead and zinc from a Pb-Zn oxidizing roasting cinder. The oxidizing roasting cinder—containing 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S—was obtained from a Pb-Zn sulfide in the Hanyuan area of China by a flotation-rotary kiln oxidizing roasting process. and lead were the main Pb- minerals, while zinc sulfate, and zinc ferrite were the main Zn-bearing minerals. The results show that a part of lead contained in lead oxide is transformed to anglesite, and a 3PbO PbSO H O-dominated new lead mineral after acid leaching. A zinc · 4· 2 leaching efficiency of 96.7% was obtained under the leaching conditions used: a leaching temperature of 55 ◦C; a leaching time of 90 min; a dosage of 20%; a sulfurous acid dosage of 4%; a cinder particle size of <0.3 mm; and a solid-liquid ratio of R = 1:4. After the gravity shaking-table separation, a lead concentrate with 50.2% Pb, 2.33% Zn and lead recovery of 86.0% was produced. The main chemical compounds in leaching residue are anglesite, 3PbO PbSO H O, SiO and ZnFe O , · 4· 2 2 2 4 while the main chemical compounds in lead concentrate are anglesite, 3PbO PbSO H O and SiO . · 4· 2 2 Keywords: lead; zinc; oxidizing roasting cinder; acid leaching; gravity separation

1. Introduction Lead and zinc mineral resources in China are abundant and widely distributed, but relatively concentrated. According to the statistical data of the United States geological survey, the world’s confirmed reserves of lead were 89 million tons in 2013, of which China’s reserves were 14 million tons, accounting for 15.7% of the total. The global reserves of zinc were 250 million tons, of which China’s reserves were 43 million tons, accounting for 17.2% of the total. Chinese reserves of both metals are ranked second in the world. The mineral resources report of the Ministry of Land and Resources shows that the confirmed reserves of lead and zinc were 73.849 million tons and 144.861 million tons in China, respectively. According to the degree of ore oxidation, lead-zinc ore can be divided into sulfide ore (lead or zinc oxidation < 10%), mixed ore (lead or zinc oxidation 10–30%) and oxidized ore (lead or zinc oxidation > 30%). The degree of ore oxidation is closely related to the selection of the separation method and technological process used for recovery. Sulfide ore accounts for the majority of lead-zinc ore in China, as as worldwide. According to the differences in mineral floatability and floating speed, the process for lead-zinc ore can be divided into three categories: selective flotation; bulk flotation; and combined flow. On this basis, according to specific ore properties, many technological processes have

Metals 2020, 10, 465; doi:10.3390/met10040465 www.mdpi.com/journal/metals MetalsMetals2020 2020, 10,, 46510, x FOR PEER REVIEW 2 of 214 of 14

in mineral floatability and floating speed, the process for lead-zinc ore can be divided into three beencategories: derived, suchselective as the flotation; direct bulk selective flotation; flotation and combined process, partial flow. On selective this basis, flotation according process, to specific floatable flotationore properties, process, many asynchronous technological flotation processes process, have partialbeen derived, bulk flotation such as the process, direct selective fully bulk flotation flotation process,process, branch partial series selective flotation flotation process, process, gravity floatabl separation-floatatione flotation process, asynch processronous and flotation floatation-gravity process, separationpartial bulk process, flotation etc. process, Sometimes fully a bulk combined flotation process process, of branch beneficiation series fl andotation process, gravity is used to processseparation-floatation extremely refractory process bulk and lead-zinc floatation-gravity or lead-zinc separation oxide process, ores etc. [1–3 Sometimes]. a combined processThere areof beneficiation a large amount and metallurgy of Cu-Pb-Zn is used polymetallic to process sulfide extremely ores refractory in the Hanyuan bulk lead-zinc area of ores China, or of whichlead-zinc the main oxide metal-containing ores [1–3]. minerals are , , , marmatite and pyrite. The mainThere gangue are a mineralslarge amount are ,of Cu-Pb-Zn calcite, polymetallic quartz, , ores mica,in the chlorite,Hanyuan montmorillonite,area of China, etc.of Assemblages which the main of galena, metal-containing sphalerite andminerals marmatite are chalcopyrite, (Fe-sphalerite) galena, are sphalerite, currently producedmarmatite asand solid pyrite. The main gangue minerals are dolomite, calcite, quartz, feldspar, mica, chlorite, solutions with a fine grain size, and occur in these ores. Copper concentrate with more than 20% Cu, montmorillonite, etc. Assemblages of galena, sphalerite and marmatite (Fe-sphalerite) are currently lead concentrate with more than 62% Pb and mixed Pb-Zn sulfide concentrate with about 40% Pb + produced as solid solutions with a fine grain size, and occur in these ores. Copper concentrate with Zn weremore obtainedthan 20% byCu, floatation lead concentrate in the localwith dressingmore than plant. 62% Pb The and elements mixed Pb-Zn associated sulfide with concentrate the sulfide mineralswith about are sulfur, 40% ,Pb+Zn ,were obtained , by floatati ,on in the andlocal . dressing plant. The separation The elements of lead andassociated zinc from with the lead-zinc the sulfide sulfide minerals ore isare di ffisulfur,cult usingsilver, just cadmium, flotation, gallium, due to theindium, close arsenic association and of leadantimony. with zinc. The With separation the development of lead and of zinc the market,from the thelead-zinc ofsulfide lead-zinc ore is mixed difficult concentrate using just has tendedflotation, to decrease, due to the while close a association large number of lead of with these zinc. concentrates With the development are unsalable, of the neither market, of the which price are conduciveof lead-zinc to the mixed improvement concentrate of economichas tended benefits. to decrease, Therefore, while a it large is crucial number that of the these effective concentrates separation of leadare unsalable, and zinc mixedneither lead-zinc of which are concentrate conducive is to studied the improvement [4–9]. of economic benefits. Therefore, it isAt crucial present, that sulfatingthe effective roasting-leaching separation of lead and and high-pressure zinc mixed lead-zinc acid leaching-extraction concentrate is studied separation [4–9]. are the mainAt metallurgical present, sulfating methods roasting-leaching to process Pb-Znand high-pressure bulk concentrate. acid leaching-extraction Zinc sulfate is separation obtained byare the leachingthe main and metallurgical formation of methods line to process products, Pb-Zn while bulk the concentrate. lead sulfate Zinc remains sulfate in is the obtained leaching by residue. the Theleaching high-pressure and formation acid leaching of crystal process line products, is used while to dissolve the lead sulfate lead and remains zinc in into thePb leaching2+and residue. Zn2+ , The high-pressure acid leaching process is used to dissolve lead and zinc into Pb2+and Zn2+ ions, respectively. Then, lead and zinc are further extracted by electrodeposition and other processes to respectively. Then, lead and zinc are further extracted by electrodeposition and other processes to achieve effective separation [10–16]. In the local enterprises, lead-zinc bulk concentrate is achieve effective separation [10–16]. In the local mining enterprises, lead-zinc bulk concentrate is treatedtreated in a in rotary a rotary kiln, kiln, and and the the schematic schematic diagram diagram isis shownshown in in Figure Figure 1.1.

E Sensor D Sensor C Sensor B Sensor A Sensor

Feeding Combustion chamber hole Φ3.6 m Discharge hole

2 m 4.0 m 5.5 m 6.0 m 2.5 m

FigureFigure 1. 1.Schematic Schematic diagramdiagram of rotary kiln kiln in in the the local local factory. factory.

TheThe roasting roasting temperature temperature and and roasting roasting atmosphereatmosphere are are the the two two key key factors, factors, and and are are the the most most difficult to control in the oxidizing roasting process of mixed lead-zinc concentrate. The amount of difficult to control in the oxidizing roasting process of mixed lead-zinc concentrate. The amount of air air in and out of the blower is used to control the roasting atmosphere. If the content is in and out of the blower is used to control the roasting atmosphere. If the oxygen content is sufficient, sufficient, the roasting products are mainly sulfates. If the oxygen content is insufficient, the roasting the roasting products are mainly sulfates. If the oxygen content is insufficient, the roasting products products are mainly . It is difficult to accurately control the roasting temperature of a rotary are mainlykiln, which oxides. is generally It is diffi cultcontrolled to accurately by temperature control the installed roasting sensors. temperature The rotary of a rotarykiln (B-C) kiln, region which is generallyrepresents controlled the high by temperaturetemperature installedbelt, while sensors. the rotary The rotarykiln (C-D) kiln region (B-C)region represents represents the middle the high temperaturetemperature belt, zone while and the the rotary rotary kilnkiln (C-D)(D-E) region region represents represents the the low middle temperature temperature zone. The zone mixed and the rotary kiln (D-E) region represents the low temperature zone. The mixed lead-zinc concentrate can react with oxygen in the middle and high temperate zones, so the local factory can only control the roasting temperature to fluctuate within a certain range in the production process, resulting in the mixture of Metals 2020, 10, 465 3 of 14 metal oxide and metal sulfate in the oxidized roasting cinder. The separation of lead and zinc from oxidized calcined by water leaching and the spiral chute gravity separation process was not ideal. At present, more than 60% of lead and over 70% of zinc can be recovered by oxidizing roasting, water leaching and spiral chute gravity separation from the mixed Pb-Zn sulfide concentrate. However, as the atmosphere is difficult to control during rotary kiln roasting, some lead oxide, zinc oxide and zinc ferrite mineral phases still exist in the oxidizing roasting cinders, resulting in the poor separation of lead and zinc. Therefore, separation tests of lead and zinc from the oxidizing roasting cinder produced from the mixed Pb-Zn sulfide concentrate were carried out. From this study, a technical scheme for the resource utilization of lead-zinc sulfide ore resources in the Hanyuan area of China was developed.

2. Materials and Methods

2.1. Sample Characterization The main sulfide minerals identified are galena, sphalerite, marmatite and pyrite; while the main gangue minerals in the mixed Pb-Zn sulfide concentrate are calcite, dolomite, dolomite, mica, chlorite, feldspar, etc. Galena reacts with oxygen to form anglesite (PbSO4) and lead oxide (PbO), while sphalerite reacts with oxygen to form zinc sulfate (ZnSO4) and zinc oxide (ZnO), and marmatite reacts with oxygen to form zinc sulfate (ZnSO4), zinc oxide (ZnO) and zinc ferrite (Zn2FeO4). Pyrite reacts with oxygen to form (Fe2O3) or (Fe3O4). The main chemical reactions that may occur in the sulfuric acid leaching system are shown in Equations (1)–(16), and these are confirmed by characterization results, as described below.

2PbS + 3O 2PbO + 2SO (1) (s) 2(g) → (s) 2(g) PbO + SO + 0.5O PbSO (2) (s) 2(g) 2(g) → 4(s) 2ZnS + 3O 2ZnO + 2SO (3) (s) 2(g) → (s) 2(g) ZnO + SO + 1/2O ZnSO (4) (s) 2(g) 2(g) → 4(s) (Zn,Fe)S + 2O ZnO + PbO + SO (5) (s) 2(g) → (s) (s) 2(g) 4FeS + 7O 2Fe O + 4SO (6) 2(s) 2(g) → 2 3(s) 2(g) 3FeS + 8O Fe O + 6SO (7) 2(s) 2(g) → 3 4(s) 2(g) ZnO + Fe O ZnFe O (8) (s) 2 3(s) → 2 4(s) CaCO CaO + CO (9) 3(s) → (s) 3(g) CaMg(CO ) CaO + MgO + 2CO (10) 3 2(s) → (s) (s) 2(g) CaO + Fe O CaFe O (11) (s) 2 3(s) → 2 4(s) MgO + Fe O MgFe O (12) (s) 2 3(s) → 2 4(s) CaO + SiO CaO SiO (13) (s) 2(s) → · 2(s) MgO + SiO MgO SiO (14) (s) 2(s) → · 2(s) CaO + Al O CaO Al O (15) (s) 2 3(s) → · 2 3(s) MgO + Al O MgO Al O (16) (s) 2 3(s) → · 2 3(s) The sample was collected from the Pb-Zn sulfide ore dressing plant by the flotation-oxidizing roasting process. The oxidizing roasting cinders contained 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S. Anglesite (PbSO4) and lead oxide (PbO) were the main Pb-bearing minerals, while zinc sulfate (ZnSO4), zinc oxide (ZnO) and zinc ferrite (Zn2FeO4) were the main Zn-bearing minerals. Gangue minerals were Metals 2020, 10, 465 4 of 14

quartz (SiO2), ferrite (CaFe2O4), ferrite (MgFe2O4), calcium aluminate (Ca2Al2O4), etc. The chemical composition analysis of the sample is shown in Table1. The lead and zinc mineral Metals 2020, 10, x FOR PEER REVIEW 4 of 14 phase analyses for the main elements are shown in Tables2 and3, respectively, while a size fraction is shown in Table4, and the X-ray di ffraction (XRD, X Pert pro, Panaco, the Netherlands) analysis is aluminate (Ca2Al2O4), etc. The chemical composition analysis of the sample is shown in Table 1. The shownlead and in Figurezinc mineral2. phase analyses for the main elements are shown in Tables 2 and 3, respectively, while a size fraction is shown in Table 4, and the X-ray diffraction (XRD, X Pert pro, Panaco, the Table 1. Main chemical composition of the oxidizing roasting cinder (wt %). Netherlands) analysis is shown in Figure 2.

Pb Zn Fe S As CaO SiO2 Al2O3 MgO Na2O 16.9Table 30.5 1. Main 3.36 chemical 15.1 composition 0.01 3.22of the oxidizing 8.36 roasting 4.23 cinder 2.89 (wt %). 1.66

Pb Zn Fe S As CaO SiO2 Al2O3 MgO Na2O 16.9Table 30.5 2. Lead3.36 phase 15.1 analysis 0.01 of the 3.22 oxidizing 8.36 roasting 4.23 cinder 2.89 (wt %). 1.66 Lead in Lead Sulfate Lead in Lead Oxide Lead in Lead Sulfide Lead in 10.3Table 2. Lead phase 6.52analysis of the oxidizing 0.02 roasting cinder (wt %). 0.06 Lead in Lead Sulfate Lead in Lead Oxide Lead in Lead Sulfide Lead in Lead Carbonate 10.3 Table 3. Zinc phase6.52 analysis of the oxidizing0.02 roasting cinder (wt %). 0.06

Zinc in Zinc Sulfate Zinc in Zinc Oxide Zinc in Lead Sulfide Zinc in Zinc Carbonate Table 3. Zinc phase analysis of the oxidizing roasting cinder (wt %). 20.2 10.2 0.06 0.04 Zinc in Zinc Sulfate Zinc in Zinc Oxide Zinc in Lead Sulfide Zinc in Zinc Carbonate 20.2 Table 4. Size fraction10.2 analysis of the oxidizing 0.06 roasting cinder. 0.04

Grade (wt %) Distribution (wt %) Fractions (mm)Table 4. Size Content fraction (wt analysis %) of the oxidizing roasting cinder. Pb Zn Pb Zn Grade (wt %) Distribution (wt %) Fractions2.00~ +(mm)1.00 Content 25.7 (wt %) 15.4 28.7 23.4 24.1 − Pb Zn Pb Zn 1.00~+0.50 25.1 17.8 31.2 26.5 25.8 −2.00~+1.00− 25.7 15.4 28.7 23.4 24.1 0.50~+0.25 16.6 18.3 27.9 18.0 15.2 −1.00~+0.50− 25.1 17.8 31.2 26.5 25.8 −0.50~+0.250.25~+0.15 16.6 11.8 15.918.3 27.9 30.6 18.0 11.0 11.815.2 − −0.25~+0.150.15~+0.10 11.8 15.9 16.915.9 30.6 34.4 11.0 16.0 18.011.8 − −0.15~+0.100.10~0 15.9 4.90 17.316.9 34.4 31.8 16.0 5.10 5.1018.0 −0.10~0− 4.90 17.3 31.8 5.10 5.10 Totals 100 16.9 30.5 100 100 Totals 100 16.9 30.5 100 100

1-PbSO4 1 2 2-ZnSO4 3-PbO 4-ZnO

5-ZnFe2O4

6-CaFe2O4

7-MgFe2O4

8-CaAl2O4 3 9-SiO2 4

Intensity (a.u.) Intensity 3 4 9 5 6 7 5 8 86 7

0 5 10 15 20 25 30 35 40 45 50 55 60 65 70 2θ (°)

FigureFigure 2.2. XRD didiffractogramffractogram and identified identified phases of of the the oxidizing oxidizing roasting roasting cinder. cinder. 2.2. Chemical Reagent and Equipment 2.2. Chemical Reagent and Equipment The main chemical reagents used in this test are sulfuric acid (purity 98%) and sulfurous acid (purity The main chemical reagents used in this test are sulfuric acid (purity 98%) and sulfurous acid 99.5%), both obtained by Chengdu Jinshan Chemical Reagent Co., Ltd. (Chengdu, China). The main (purity 99.5%), both obtained by Chengdu Jinshan Chemical Reagent Co., Ltd (Chendu, China). The 3 equipmentmain equipment used for used the for experiment the experime wasnt a magneticwas a magnetic stirrer (1.5stirrer dm (1.5, Jilin dm Exploration3, Jilin Exploration Machinery Machinery Factory, Factory, Changchun, China), a mechanical stirrer (10 dm3, Jilin Exploration Machinery Factory, Changchun, China), a cone (XMQ—Ф240 × 90 mm, Jilin Exploration Machinery Factory, Changchun, China), a disc grinder (Ф300 × 150 mm, Jilin Exploration Machinery Factory, Changchun, China), a gravity shaking-table (LY-2100 × 1050 mm, Sichuan Jiangyou Province Mining Machinery

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Changchun, China), a mechanical stirrer (10 dm3, Jilin Exploration Machinery Factory, Changchun, China), a cone ball mill (XMQ—Φ240 90 mm, Jilin Exploration Machinery Factory, Changchun, China), × a disc grinder (Φ300 150 mm, Jilin Exploration Machinery Factory, Changchun, China), a gravity × shaking-table (LY-2100 1050 mm, Sichuan Jiangyou Province Mining Machinery Co., Ltd., Jiangyou, × China), a drying box (Shanghai Shiyan Electric Furnace Co., Ltd., Shanghai, China), a vacuum filter (Φ300, Southwest Chengdu Experimental Equipment Co., Ltd., Chengdu, China), a ceramic mortar (Φ100, Beijing Grinder Instrument Co., Ltd., Beijing, China) and a separating funnel (100 mL, 500 mL, 1000 mL, Taizhou Gaogang Kejia Experimental Equipment Co., Ltd., Taizhou, China).

Metals 2020, 10, x FOR PEER REVIEW 5 of 14 2.3. Experimental Procedure Co., Ltd, Jiangyou, China), a drying box (Shanghai Shiyan Electric Furnace Co., Ltd, Shanghai, SulfuricChina), acid a vacuum leaching filter and (Ф gravity300, Southwest shaking-table Chengdu Experiment separational wereEquipment used Co., to separateLtd, Chengdu, lead and zinc from the oxidizingChina), a ceramic roasting mortar cinder, (Ф100, Beijing and this Grinder process Instrument is shown Co., Ltd, in Beijing, Figure China)3. To and conduct a separating acid leaching, raw materialsfunnel (a (100 mass mL, of 500 100 mL, g for1000 each mL, test)Taizhou are Gaogang mixed withKejia theExperimental sulfuric acidEquipment in a 5–30%Co., Ltd, proportion Taizhou, China). and placed into the stirrer. Water was added to adjust the solid:liquid ratio to the target value, and the sulfuric2.3. acidExperimental reacts Procedure with the oxidizing roasting cinder to generate insoluble PbSO4 and soluble ZnSO4. AfterSulfuric solid-liquid acid leaching separation, and gravity ZnSO shaking-table4 reports to se theparation lixivium were used and to PbSO separate4 to lead the and leaching zinc residue. Zinc-bearingfrom lixiviumthe oxidizing is usedroasting for cinder, the preparationand this process of is ZnSO shown in7H FigureO. Sulfuric3. To conduct acid acid leaching leaching, residue (a 4· 2 mass of 500raw g materials for each (a test)mass andof 100 water g for each were test) mixed are mixed in a with 30–40% the sulfuric portion acid andin a 5–30% gravity proportion shaking-table to recover lead.and placed The contentsinto the stirrer. of lead Water and was zincadded in to the adju sulfuricst the solid:liquid acid leaching ratio to the residue target value, were and analyzed to the sulfuric acid reacts with the oxidizing roasting cinder to generate insoluble PbSO4 and soluble calculate the leaching efficiency for zinc, and evaluate the separation effect of zinc and enrichment ZnSO4. After solid-liquid separation, ZnSO4 reports to the lixivium and PbSO4 to the leaching residue. of lead. TheZinc-bearing contents lixivium of lead is used in the for gravitythe preparation shaking-table of ZnSO4·7H separation2O. Sulfuric concentrateacid leaching residue and tailings (a were analyzedmass to calculate of 500 g for the each lead test) recovery. and water The were calculation mixed in a 30–40% formula portion for and zinc gravity leaching shaking-table efficiency to is shown in Equationrecover (17), lead. and The lead contents recovery of lead is shownand zinc in in Equationthe sulfuric (18). acid leaching residue were analyzed to calculate the leaching efficiency for zinc, and evaluate the separation effect of zinc and enrichment of lead. The contents of lead in the gravity shaking-table separation concentrate and tailings were Zinc leaching efficiency = (Q α - Q θ )/Q α (17) analyzed to calculate the lead recovery. The calculation1 × formulaZn 2for× zincZn leaching1 × efficiencyZn is shown in Equation (17), and lead recovery is shown in Equation (18) Lead recovery = Q3 βPb/(Q3 βPb + Q4 θPb) (18) Zinc leaching efficiency× = (Q1 × αZn ×– Q2 × θZn)/Q1 × αZn (17) where Q1 is the weight of the oxidizingLead recovery roasting = Q3 × cinderβPb/(Q3 × (g); βPb +Q Q24 ×is θ thePb) weight of the leaching(18) residue

(g); Q3 is thewhere weight Q1 is the of weight the gravity of the oxidizing shaking-table roasting cinder separation (g); Q2 is concentrate the weight of (g);the leachingQ4 is the residue weight of the gravity shaking-table(g); Q3 is the weight separation of the gravity tailings shaking-table (g); αZn is separation content ofconcentrate the oxidizing (g); Q4 is roasting the weight cinder of the (%); θZn is gravity shaking-table separation tailings (g); αZn is content of the oxidizing roasting cinder (%); θZn is the content of leaching residue (%); βPb is content of gravity shaking-table separation concentrate (%); the content of leaching residue (%); βPb is content of gravity shaking-table separation concentrate (%); and θPb is content of gravity shaking-table separation tailings (%). and θPb is content of gravity shaking-table separation tailings (%).

Oxidizing roasting cinder

Sulfuric acid Mixing Water

Leaching

Solid-liquid separation

Leaching residue Lixivium

Water Mixing (concentration, 30-40%)

Gravity shaking-table separation

Lead concentrate Tailings

Figure 3. Scheme of the extraction of lead and zinc process from the oxidizing roasting cinder. Figure 3. Scheme of the extraction of lead and zinc process from the oxidizing roasting cinder. 2.4. Analysis and Characterization

Metals 2020, 10, 465 6 of 14

Metals 2020, 10, x FOR PEER REVIEW 6 of 14 2.4. Analysis and Characterization The chemical composition of solid materials was analyzed by a Z-2000 atomic absorption spectrophotometerThe chemical (Hitachi composition Co., ofLtd, solid Tokyo, materials Japan) was; the analyzed diffraction by grating a Z-2000 was atomic Zenier-tana absorption type, spectrophotometer (Hitachi Co., Ltd., Tokyo, Japan); the diffraction grating was Zenier-tana type, 1800 1800 lines/mm; the flash wavelength was 200 nm, the wavelength range was 190–900 nm; and the lines/mm; the flash wavelength was 200 nm, the wavelength range was 190–900 nm; and the automatic automatic peak seeking setting and spectral bandwidth were set to 1.3 nm and 2.6 nm, respectively, peak seeking setting and spectral bandwidth were set to 1.3 nm and 2.6 nm, respectively, in order to in order to analyze the chemical composition of the solid. Inductively coupled plasma mass analyze the chemical composition of the solid. Inductively coupled plasma mass spectrometry (8800 spectrometry (8800 ICP-MS/MS, Agilent Inc., Santa Clara, CA, USA) was used to ICP-MS/MS, Agilent Technologies Inc., Santa Clara, CA, USA) was used to determine the Zn, Pb and determine the Zn, Pb and Fe contents of lixivium. The phase compositions of solid substances (the Fe contents of lixivium. The phase compositions of solid substances (the oxidizing roasting cinders, oxidizing roasting cinders, leaching residue, lead concentrate and gravity shaking-table separation leaching residue, lead concentrate and gravity shaking-table separation tailings) were analyzed by tailings) were analyzed by XRD. The microstructure of the solid products was observed by SEM XRD. The microstructure of the solid products was observed by SEM (S440, Hirschmann Laborgerate (S440, Hirschmann Laborgerate GmbH & Co. KG, Eberstadt, Germany) equipped with an energy- GmbH & Co. KG, Eberstadt, Germany) equipped with an energy-dispersive X-ray spectroscopy dispersive(EDS) detector X-ray (UItra55, spectroscopy CarlzeissNTS (EDS) GmbH, detect Hirschmannor (UItra55, Laborgerate CarlzeissNTS GmbH &GmbH, Co. KG, Hirschmann Eberstadt, LaborgerateGermany). GmbH & Co. KG, Eberstadt, Germany).

3.3. Results Results and and Discussion Discussion

3.1.3.1. Separation Separation of of Zinc Zinc Using Using Sulfuric Sulfuric Acid Leaching

3.1.1.3.1.1. Sulfuric Sulfuric Acid Acid Dosage Dosage HydrochloricHydrochloric acid, acid, sulfuric sulfuric acid and nitric acidacid areare commonlycommonly used used to to extract extract lead lead and and zinc zinc by by acid leaching. If or is used as the leaching agent, PbCl2 or Pb(NO3)2 and acid leaching. If hydrochloric acid or nitric acid is used as the leaching agent, PbCl2 or Pb(NO3)2 and ZnCl2 or Zn (NO3)2 will enter into the leaching solution at the same time, which prevents the effective ZnCl2 or Zn (NO3)2 will enter into the leaching solution at the same time, which prevents the effective separation of lead and zinc. When sulfuric acid is used as the leaching agent, the products are an separation of lead and zinc. When sulfuric acid is used as the leaching agent, the products are an insoluble PbSO4 precipitate and soluble ZnSO4, and lead and zinc can be separated by solid-liquid insoluble PbSO4 precipitate and soluble ZnSO4, and lead and zinc can be separated by solid-liquid separation. The effect of using different sulfuric acid dosages is shown in Figure4a. separation. The effect of using different sulfuric acid dosages is shown in Figure 4a.

30 120 100 (a) (b) Zn Leaching efficiency 28 Zn Leaching efficiency 32 98 110 Pb grade 26 Pb grade 96 Zn grade 28 Zn grade 24 100 94

22 90 24 92 20 90 80 20 18 88

16 70 16 86 14 84 60 12 12 82

10 50 8 80 Leaching efficiency (%) efficiency Leaching

8 (%) efficiency Leaching Leaching time: 60 min 40 78 Grade in leaching residue (%) residue leaching in Grade 4 Grade in leaching residue (%) 6 Sample particle size: ＜2 mm Sulfuric acid dosage: 20% 76 30 Sample particle size: ＜2 mm 4 Leacing temperature: 25 ℃ 74 0 Leacing temperature: 25 ℃ 2 Solid-liquid ratio:R=m ∶m =1∶ 3 Solid liquid 20 Solid-liquid ratio:R=m ∶m =1∶ 3 72 -4 Solid liquid 0 70 0 5 10 15 20 25 30 20 30 40 50 60 70 80 90 100 110 120 130 Mass fraction (wt %) Time (min)

Figure 4. Effect of sulfuric acid dosage (a) and leaching time (b) on the leaching efficiency of zinc and

metal grades in the leaching residue.

FigureResults 4. inEffect Figure of sulfuric4a confirm acid thatdosage increasing (a) and leaching sulfuric acidtime dosage(b) on the was leaching conducive efficiency to improving of zinc and zinc leachingmetal e gradesfficiency in andthe leaching enriching residue. lead grade in the leaching residue. When the amount of sulfuric acid exceeded 20%, the leaching efficiency of zinc increases slightly. Considering the cost of reducing acid consumptionResults in in Figure the production 4a confirm process that increasing and the sulfuric of acid equipment dosage bywas high conducive concentration to improving acid, a zincsulfuric leaching acid dosageefficiency of 20%and enriching was suitable. lead The grade leaching in the e leachingfficiency residue. of zinc was When 79.2%, the andamount the grades of sulfuric of acidlead exceeded and zinc in20%, leaching the leaching residue efficiency were 25.5% of andzinc 8.86%,increases respectively. slightly. Considering the cost of reducing acid consumption in the production process and the corrosion of equipment by high concentration acid, a sulfuric acid dosage of 20% was suitable. The leaching efficiency of zinc was 79.2%, and the grades of lead and zinc in leaching residue were 25.5% and 8.86%, respectively.

3.1.2. Leaching Time The test results for different leaching times are shown in Figure 4b. Prolonging the leaching time is conducive to the dissolution of zinc, and the leaching reaction is more complete. However, when

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3.1.2. Leaching Time Metals 2020, 10, x FOR PEER REVIEW 7 of 14 The test results for different leaching times are shown in Figure4b. Prolonging the leaching time is conducive to the dissolution of zinc, and the leaching reaction is more complete. However, the leaching time is too long, it becomes easier to dissolve non-target elements in the sample. This when the leaching time is too long, it becomes easier to dissolve non-target elements in the sample. increases the impurity content in the leaching solution, and therefore affects the separation and This increases the impurity content in the leaching solution, and therefore affects the separation and purification of the zinc in the leaching solution. When the leaching time was extended to 90 min, the purification of the zinc in the leaching solution. When the leaching time was extended to 90 min, leaching efficiency of zinc increased slightly. Compared with the leaching time of 120 min, the the leaching efficiency of zinc increased slightly. Compared with the leaching time of 120 min, the leaching efficiency of zinc increased from 84.8% to 85.5%. The main reason for this small change is leaching efficiency of zinc increased from 84.8% to 85.5%. The main reason for this small change is thatthat as as the the reaction reaction time increases, thethe acidacid inin thethe system system is is consumed, consumed, and and there there is is little little impetus impetus for for thethe reaction reaction to to continue. Therefore, aa leachingleaching timetime ofof 90 90 min min was was selected selected as as most most appropriate, appropriate, the the leachingleaching efficiency efficiency of zinc was 84.8%,84.8%, andand thethe gradesgradesof of lead lead and and zinc zinc in in the the leaching leaching residue residue were were 25.7%25.7% and and 7.08%, 7.08%, respectively.

3.1.3.3.1.3. Leaching Leaching Temperature Temperature TheThe effects effects of didifferentfferent leachingleaching temperaturestemperatures are are shown shown in in Figure Figure5a. 5a. The The results results show show that that temperaturetemperature has has a a significant significant influenceinfluence onon thethe leaching leaching e ffiefficiency.ciency. When When the the leaching leaching temperature temperature roserose to to 55 55 °C,◦C, the leaching eefficiencyfficiency ofof zinczinc waswas 89.1%; 89.1%; when when the the leaching leaching temperature temperature was was further further increasedincreased toto 7575◦ C,°C, the the increase increase of theof leachingthe leaching efficiency efficiency of zinc of was zinc relatively was relatively small. With small. the increaseWith the increaseof temperature, of temperature, the influence the influence of sulfate of ions sulfate in the ions solution in the on solution minerals on increases, minerals which increases, is beneficial which is beneficialto the leaching to the of leaching valuable of metal valuable zinc. metal When zinc. the temperatureWhen the temperature reached 55 ◦reachedC, the leaching 55 °C, the effi leachingciency efficiencyof other elements of other will elements increase will significantly increase assignif theicantly temperature as the continues temperature to increase. continues Therefore, to increase. the Therefore,leaching temperature the leaching was temperature 55 ◦C, the leaching was 55 e ffi°C,ciency the le ofaching zinc was efficiency 89.2%, the of leadzinc grade was was89.2%, increased the lead gradeto 28.8%, was and increased the zinc to grade 28.8%, was and reduced the zinc to grade 5.66% was in the reduced leaching to residue. 5.66% in the leaching residue.

40 100 100 (a) Zn Leaching efficiency (b) 98 32 35 Pb grade 96 Zn grade 28 95 30 94

92 24 25 90 90 20 20 88

86 16 15 Zn Leaching efficiency 85 84 12 Pb grade 10 82 Zn grade 80 8 80

5 (%) efficiency Leaching 78 Leaching efficiency (%) Grade in leaching residue (%) residue leaching in Grade Grade in leaching residue (%) residue leaching in Grade Sulfuric acid dosage: 20% 4 Sulfuric acid dosage: 20% 0 76 Leaching time: 90 min Sample particle size: ＜2 mm 75 Leacing time: 90 min 74 0 Leaching temperature: 55 ℃ -5 Solid-liqiuid ratio: R=m ∶m =1∶ 3 Solid-liquid ratio:R=mSolid∶mliquid =1∶ 3 72 solid liquid -4 -10 70 70 20 25 30 35 40 45 50 55 60 65 70 75 80 -2 -1.5 -1 -0.5 -0.3 -0.15 -0.075 Temperature (℃) Fineness (mm)

Figure 5. Effect of leaching temperature (a) and fineness (b) on the leaching efficiency of zinc and the metal grades in the leaching residue.

3.1.4.Figure Fineness 5. Effect of the of leaching Oxidizing temperature Roasting Cinder(a) and fineness (b) on the leaching efficiency of zinc and the metal grades in the leaching residue. The effects of different particle size fractions are shown in Figure5b. Since the mixed lead-zinc concentrate was oxidized in a rotary kiln, zinc sulfate and lead sulfate are both present in the roasting 3.1.4. Fineness of the Oxidizing Roasting Cinder cinder. If wet grinding is adopted to change the material size, it is not conducive to the control of the liquid-solidThe effects ratio of in different subsequent particle leaching size testsfractions and theare concentrationshown in Figure of Zn 5b.2+ Sincein the the leaching mixed solution. lead-zinc concentrateFurthermore, was it isoxidized not conducive in a rotary to the kiln, subsequent zinc sulfate separation and lead or sulfate the recovery are both of present zinc in in the the leaching roasting cinder.solution. If wet The resultsgrinding show is adopted that decreasing to change the the particle material size of size, the it oxidizing is not conducive roasting residue to the control is beneficial of the liquid-solidto increase the ratio leaching in subsequent efficiency leaching of zinc and tests the and lead the grade concentration in leaching residue.of Zn2+ in However, the leaching the decrease solution. Furthermore,of feed size also it is means not conducive an increase to inthe the subsequent grinding cost. separation As the or size the of recovery the roasting of zinc ore decreases,in the leaching its solution.specific surface The results area increases, show that and decreasing mineral liberation the particle is enhanced. size of Both the of oxidizing these changes roasting are conducive residue is beneficial to increase the leaching efficiency of zinc and the lead grade in leaching residue. However, the decrease of feed size also means an increase in the grinding cost. As the size of the roasting ore decreases, its specific surface area increases, and mineral liberation is enhanced. Both of these changes are conducive to the leaching reaction, so the leaching efficiency also increases gradually. When the particle size of leaching materials was <0.3 mm, the leaching efficiency of zinc was 90.1%, the lead grade was enriched to 30.5%, and the zinc grade was reduced to 5.12% in leaching residue.

Metals 2020, 10, 465 8 of 14

to the leaching reaction, so the leaching efficiency also increases gradually. When the particle size of leaching materials was <0.3 mm, the leaching efficiency of zinc was 90.1%, the lead grade was enriched Metalsto 30.5%, 2020, 10 and, x FOR the PEER zinc gradeREVIEW was reduced to 5.12% in leaching residue. 8 of 14

3.1.5.3.1.5. Solid-Liquid Solid-Liquid Ratio Ratio (R (R == mmsolidsolid:m:liquidmliquid) ) TheThe test test results results of of different different solid-liquid solid-liquid ratios are shown inin FigureFigure6 a.6a. Once Once the the solid-liquid solid-liquid ratio ratio reachedreached 1:4, 1:4, the the viscosity viscosity of of the the leaching leaching solution solution changed little.little. Hence, didiffusionffusion ratesrates of of species species to to andand from from the the reacting reacting mineral mineral surfaces surfaces did did not not increa increasese significantlysignificantly withwith thethe furtherfurther decrease decrease of of the the solid-liquidsolid-liquid ratio. ratio. Hence, Hence, the the leaching leaching efficiency efficiency increased increased slowly. AA solid-liquidsolid-liquid ratio ratio for for the the leaching leaching ofof RR == 1:41:4 waswas thereforetherefore suitable. suitable. The The leaching leaching efficiency efficiency of zinc of was zinc 91.4%, was the91.4%, lead the grade lead was grade enriched was enrichedto 32.8%, to and 32.8%, the zincand gradethe zinc was grade reduced was toreduced 5.08% into leaching5.08% in residue. leaching residue.

100 40 100 (b) 32 (a) 98 35 98 28 Zn Leaching efficiency 30 Pb grade 96 96 24 Zn grade 25 94 94 20 20 Zn Leaching efficiency 92 Pb grade 92 16 15 Zn grade 12 90 10 90 5 8 88 88 0 4 86 Sulfuric acid dosage: 20% 86 Leaching efficiency (%) efficiency Leaching Sulfuric acid dosage: 20% -5 Leaching time: 60 min Leaching efficiency (%) Grade in leaching residue (%) residue in leaching Grade 0 Leaching time: 90 min 84 Grade in leaching residue (%) Sample particle size: ＜0.3 mm 84 -10 -4 Sample particle size: ＜0.3 mm Leaching temperature: 55 ℃ Leaching temperature: 55 ℃ 82 -15 82 Solid-liquid ratio: R=mSolid∶mliquid =1∶ 4 -8 80 -20 80 11∶ 12∶ 13∶ 14∶ 15∶ 16∶ 17∶ -202468101214 R=m ∶m (g/g) solid liquid Mass fraction (wt %)

Figure 6. Effect of the solid-liquid ratio (a) and sulfurous acid dosage (b) on the leaching efficiency of zinc and the metal grades in the leaching residue.

3.1.6.Figure Additive 6. Effect Dosage of the solid-liquid ratio (a) and sulfurous acid dosage (b) on the leaching efficiency of zincThe and eff ectsthe metal of di ffgradeserent sulfurousin the leaching acid residue. dosage are shown in Figure6b. The mineral analysis of the sample indicated the presence of ZnFe2O4, which does not readily react with sulfuric acid to produce 3.1.6. Additive Dosage ZnSO4 and Fe2(SO4)3 during the leaching process. Combining this with the previous relevant study, theThe addition effects of oxidizerof different in the sulfurous leaching acid process dosage was ar beneficiale shown to in improve Figure the6b. leachingThe mineral efficiency. analysis The of thepredominance sample indicated diagram the of presenceϕ-pH for of the ZnFe SO2-ZnFe2O4, which2O4-H 2doesO system not readily at different react temperatures with sulfuric is shownacid to producein Figure ZnSO7. There4 and are Fe2 two(SO4 mechanisms)3 during the forleaching the decomposition process. Combining of ZnFe this2O 4withunder the acidic previous conditions. relevant 2+ study,Under the the addition oxidizing of conditions, oxidizer thein firstthe leaching step of ZnFe process2O4 decomposition was beneficial is intoto improve Fe2O3 and the Zn leaching, and 3+ efficiency.the second The step predominance is Fe dissolution diagram from of Fe 2φO-pH3. Under for the the SO reducing2-ZnFe2 conditions,O4-H2O system after ZnFeat different2O4 is 2+ 2+ temperaturesdecomposed is into shown Fe2O 3inand Figure Zn 7., Fe There2O3 is ar reductivelye two mechanisms dissolved for to Fethe .decomposition When the of potential ZnFe2O4 2+ underwas controlledacidic conditions. in the (B) Under region, the zinc oxidizing in ZnFe conditions,2O4 entered the the first solution step of in ZnFe the form2O4 decomposition of Zn , while is remained in2+ the residue in the form of3+ Fe O . When the redox potential was controlled in the into Fe2O3 and Zn , and the second step is Fe dissolution2 3 from Fe2O3. Under the reducing conditions, (D), (E) and (F) regions, zinc and iron in ZnFe O entered the solution in the form of Zn2+ and Fe2+, after ZnFe2O4 is decomposed into Fe2O3 and Zn2 2+4 , Fe2O3 is reductively dissolved to Fe2+. When the respectively [17–20]. redox potential was controlled in the (B) region, zinc in ZnFe2O4 entered the solution in the form of The results in Figure5b show that increasing the appropriate amount of sulfurous acid is conducive Zn2+, while iron remained in the residue in the form of Fe2O3. When the redox potential was controlled to improving the extraction efficiency of zinc. When the amount of sulfurous acid exceeded 4%, in the (D), (E) and (F) regions, zinc and iron in ZnFe2O4 entered the solution in the form of Zn2+ and the leaching efficiency of zinc changes little. These results indicated that sulfurous acid can play Fe2+, respectively [17–20]. a synergistic role in enhancing the leaching efficiency of zinc using sulfuric acid. A 4% dosage of sulfurous acid was deemed sufficient, and allowed the redox potential to be controlled within the (D), and with a zinc leaching efficiency of 96.5%. The lead grade was enriched to 33.6% and the zinc grade was reduced to 2.12% in the leaching residue.

Metals 2020, 10, 465 9 of 14 Metals 2020, 10, x FOR PEER REVIEW 9 of 14

0246810 1.2 10 O2 Zn2+ 3+ (A) Fe Zn2+ H2O Fe2O3 8 0.8 (B) ZnFe2O4

2+ 6 Zn (C) Fe2+ (D) /V

ϕ 0.4

4 H+

0.0 H 2 2 (E) (F) H2SO3 - HSO3 -0.4 0 -2-10123456 pH

FigureFigure 7.7. PredominancePredominance diagram diagram of of φϕ-pH-pH for for SO SO2-ZnFe2-ZnFe2O24-HO42-HO 2systemO system (25 °C—Full (25 ◦C—Full lines, lines, 100 °C— 100 ◦DashedC—Dashed lines). lines). 3.1.7. Repeated Scale-up Test The results in Figure 5b show that increasing the appropriate amount of sulfurous acid is conduciveIn order to to improving further validate the extraction test repeatability, efficiency repeated of zinc. scale-up When teststhe wereamount carried of sulfurous out under acid the conditionsexceeded 4%, employed the leaching as follows: efficiency a sulfuric of zinc acid change dosages little. of 20%;These a results leaching indicated time of 60that min; sulfurous a leaching acid temperaturecan play a synergistic of 55 ◦C; arole cinder in enhancing particle size the of leaching<0.3 mm; efficiency and a solid-liquid of zinc using ratio sulfuric of R = acid.1:3. TheseA 4% testsdosage scaled of sulfurous to the ore acid mass was up deemed to 3000 gsufficient, for each test,and allowed and the resultsthe redox are potential shown in to Table be controlled5. These resultswithin showthe (D), that and the indexwith a of zinc repeated leaching experiments efficiency is of superior 96.5%. toThe the lead conditions grade was of results. enriched Lead to enters33.6% intoand thethe leachingzinc grade residue, was reduced and is enrichedto 2.12% in the leaching residueresidue. from 16.9% to 33.8%. The impurity elements also report to the residue at significant concentrations (Table6). 3.1.7. Repeated Scale-up Test Table 5. Repeated scale-up test results (wt %). In order to further validate test repeatability, repeated scale-up tests were carried out under the conditions employedContents in as the follows: Oxidizing a sulfuric Roasting acid Cinder dosage Contents of 20%; in Leachinga leaching Residue time of 60 min; a leaching Repeat Leaching temperature of 55 °C;Pb a cinder particle size Zn of <0.3 mm; and Pb a solid-liquid Zn ratio of RE ffi= ciency1:3. These of Zinc tests scaled1 to the ore mass 16.8 up to 3000 g for each 30.6 test, and the 33.9 results are shown 2.08 in Table 5. These 96.7 results show2 that the index 17.1 of repeated experiments 30.4 is superior to 33.6 the conditions 2.12 of results. Lead 96.6 enters into the leaching residue, and is enriched in the leaching residue from 16.9% to 33.8%. The impurity 3 16.9 30.4 33.9 2.05 97.1 elements also report to the residue at significant concentrations (Table 6). 4 17.0 30.6 33.7 2.11 96.5

5 16.7Table 5. Repeated 30.5 scale-up test 33.9 results (wt %). 2.06 96.7 6 16.8 30.4 33.7 2.07 96.7 Contents in the Oxidizing Contents in Leaching Average 16.9 30.5 33.8 2.08Leaching Efficiency 96.7 of Repeat Roasting Cinder Residue Range (R = Emax Emin) = 0.60 Zinc Pb Zn −Pb Zn PN di 1 16.8 30.6 33.9 i=1| | 2.08 96.7 Arithmetic mean error (δ = N ) = 96.7 2 17.1 30.4 33.6 2.12 96.6 N N 2 = P 2 = P ( ) Sum square variation SS di Ei E = 0.21 3 16.9 30.4 i= 33.91 i=1 − 2.05 97.1 4 17.0 30.6 33.7 SS 2.11 96.5 Average deviation MS = f = 0.04 5 16.7 30.5 r 33.9 2.06 96.7 P 2 (Ei E) √ 6 16.8 Standard 30.4 deviation s = 33.7N −1 = MS 2.07= 0.20 96.7 Average 16.9 30.5 33.8− 2.08 96.7 =− Range ( REmax E min ) = 0.60 N

 di = Arithmetic mean error ( δ = i 1 ) = 96.7 N

Metals 2020, 10, x FOR PEER REVIEW 10 of 14

NN ==22 − Sum square variation SS dii() E E = 0.21 ii==11 SS Average deviation MS = = 0.04 f − 2 Metals 2020, 10, 465 ()EEi 10 of 14 Standard deviation sMS== = 0.20 N −1

Table 6. Average main element analysis of leaching residues (wt %). Table 6. Average main element analysis of leaching residues (wt %).

Pb Zn Fe S CaO SiO2 Al2O3 MgO Pb Zn Fe S CaO SiO2 Al2O3 MgO 33.833.8 2.07 2.07 0.260.26 5.12 5.12 9.18 9.18 36.9 36.9 0.25 0.25 0.45 0.45

3.2.3.2. RecoveringRecovering LeadLead fromfrom LeachingLeaching ResidueResidue byby TableTable Gravity Gravity Separation Separation GravityGravity shaking-tableshaking-table separationseparation waswas adoptedadopted toto furtherfurther recoverrecover leadlead fromfrom thethe acidacid leachingleaching residue,residue, and and the the test test procedure procedure is shownis shown in Figurein Figure8. The 8. The test resultstest results are shown are shown in Table in 7Table, and the7, and main the elementmain element analysis analysis results forresults the leadfor concentratethe lead concen generatedtrate generated are shown are in Tableshown8. Thein Table lead concentrate 8. The lead withconcentrate a lead grade with a of lead 50.2% grade and of recovery 50.2% and of 86.0%recovery was of obtained86.0% was by obtained two stages by oftwo grinding stages of and grinding three stagesand three of gravity stages shaking-table of gravity shaking-table separation. Accordingseparation. to According the quality to standard the quality for lead standard concentrate for lead in China,concentrate the lead in China, concentrate the lead meets concentrate the requirement meets the for requirement level four lead for concentrate,level four lead with concentrate, a lead content with ofa lead more content than 50%. of more than 50%.

Leaching residue

Grinding <0.154 mm occupying 90%

Gravity shaking-table separation

Gravity shaking-table separation

Grinding <0.1 mm occupying 90%

Gravity shaking-table separation

Lead concentrate Tailings

FigureFigure 8.8. SchemeScheme forfor recoveringrecovering leadlead fromfrom thethe leachingleachingresidue. residue.

TableTable 7. 7.Results Results ofof leadlead recoveryrecovery leadlead fromfrom the the leaching leaching residues residues (wt (wt %). %).

GradeGrade Recovery Recovery ProductsProductsProductivity Productivity PbPb ZnZn Pb Pb Zn Zn Lead concentrateLead concentrate 57.8 57.8 50.250.2 2.332.33 86.0 86.0 65.1 65.1 TailingsTailings 42.242.2 11.211.2 1.711.71 14.0 14.0 34.9 34.9 TotalsTotals 100 100 33.833.8 2.072.07 100100 100 100

TableTable 8. 8.Main Main elementelement analysisanalysis ofof leadlead concentrateconcentrate (wt (wt %). %).

Pb Zn Fe S CaO SiO2 Al2O3 MgO 50.2 2.36 0.11 8.12 3.59 17.62 0.16 0.11 Metals 2020, 10, x FOR PEER REVIEW 11 of 14

Pb Zn Fe S CaO SiO2 Al2O3 MgO

Metals 2020, 10, 465 50.2 2.36 0.11 8.12 3.59 17.62 0.16 0.11 11 of 14

3.3. Lead and Zinc Phase Transformation Mechanism in Sulphuric Acid Leaching Process 3.3. Lead and Zinc Phase Transformation Mechanism in Sulphuric Acid Leaching Process According to the phase of sample analysis results, the oxidizing roasting of valuable metals in ore—mainlyAccording lead to theand phase zinc ofsulfate sample and analysis oxides—exist results,s thein the oxidizing form of roasting different of valuablemineral phases, metals inthe ore—mainlyoxidizing roasting lead and cinder zinc in sulfate different and oxides—existsmineral dissolution in the mechanisms form of different for the mineral understanding phases, the of oxidizingvaluable roastingmetals such cinder as inlead, diff erentzinc and mineral iron, dissolution magnesium, mechanisms calcium and for the aluminum understanding impurities of valuable such as metalsleaching such behavior as lead, zinchave and important iron, magnesium, significance. calcium From and aluminumthe chemical impurities composition, such as leachingminerals behaviorcomposition have and important grain-sized significance. analysis From of the the complex chemical oxidizing composition, roasting minerals cinder, composition zinc was and first grain-sizedextracted by analysis sulphuric of the acid. complex The valuable oxidizing metallic roasting lead cinder, and zinc wasmainly first exist extracted in the by mineral sulphuric phases acid. of Thelead valuable sulfate, metalliclead oxide, lead zinc and zincoxide, mainly zinc existsulfate in theand mineral zinc ferrite. phases At of the lead same sulfate, time, lead there oxide, are zinc also oxide,corresponding zinc sulfate chemical and zinc reactions ferrite. Atfor thecalcium, same time,magnesium, there are aluminum also corresponding compounds chemical and sulfuric reactions acid for[21–25]. calcium, magnesium, aluminum compounds and sulfuric acid [21–25]. Therefore,Therefore, it it is is very very important important to to investigate investigate the the dissolution dissolution process process of of di differentfferent mineral mineral phases phases inin acidacid solution.solution. AccordingAccording toto thethe X-ray X-ray di diffractionffraction (XRD) (XRD) mineral mineral analysis analysis of of the the raw raw ore, ore, the the main main mineralmineral phasesphases inin thethe oxidizingoxidizing roasting roasting cinder cinder were were lead lead sulfate sulfate (PbSO (PbSO4),4), lead lead oxide oxide (PbO), (PbO), zinc zinc sulfatesulfate (ZnSO (ZnSO4),4), zinc zinc oxide oxide (ZnO), (ZnO), zinc zinc ferrite ferrite (ZnFe (ZnFe2O2O4),4), calcium calcium ferrite ferrite (CaFe (CaFe2O2O4),4), magnesium magnesium ferrite ferrite (MgFe(MgFe22OO44) andand calciumcalcium aluminatealuminate (Ca(Ca22AlAl22OO44). The main chemicalchemical reactionsreactions thatthat maymay occur occur in in the the sulfuricsulfuric acidacid leaching leaching system system were were shown shown in in Equations Equations (19)–(24). (19)–(24).

PbOPbO+(s) H+ HSO2SO4(l) →PbSO PbSO4(s)+ + HH2OO(l) (19)(19) (s) 2 4(l) → 4(s) 2 (l) ZnO(s) + H2SO4(l) → ZnSO4(l)+ H2O(l) (20) ZnO + H SO ZnSO + H O (20) (s) 2 4(l) → 4(l) 2 (l) ZnFe2O4(s) + 4H2SO3(l) + 2O2(g) → ZnSO4(l) + Fe2(SO4)3(l) + 4H2O(l) (21) ZnFe O + 4H SO + 2O ZnSO + Fe (SO ) + 4H O (21) 2 4(s) 2 3(l) 2(g) → 4(l) 2 4 3(l) 2 (l) CaFe2O4(s) + H2SO4(l) → CaSO4(s)+ 2FeSO4(l) + H2O(l) (22) CaFe O + H SO CaSO + 2FeSO + H O (22) 2 4(s) 2 4(l) → 4(s) 4(l) 2 (l) MgFe2O4(s) + 2H2SO4(l) → MgSO4(l) + 2FeSO4(l) + 2H2O(l) (23) MgFe O + 2H SO MgSO + 2FeSO + 2H O (23) 2 4(s) 2 4(l) → 4(l) 4(l) 2 (l) Ca2Al2O4(s) + 2H2SO4(l) → 2CaSO4(s) + Al2(SO4)3(l) +H2O(l) (24) Ca2Al2O4(s) + 2H2SO4(l) 2CaSO4(s) + Al2(SO4)3(l) +H2O(l) (24) XRD, scanning microscopy (SEM)→ and energy dispersive X-ray (EDS) are used to characterizeXRD, scanning the leaching electron residue microscopy and lead (SEM) concentrate, and energy and dispersive analyze the X-ray leaching (EDS) mechanism are used to of characterizesulfuric acid the system leaching solution. residue SEM-EDS and lead analysis concentrate, results and of analyze leaching the residue leaching and mechanism lead concentrate of sulfuric are acidshown system in Figure solution. 9, and SEM-EDS XRD analysis analysis results results for of leaching residue andand leadlead concentrateconcentrate are are shown shown in in FigureFigure9 ,10. and XRD analysis results for leaching residue and lead concentrate are shown in Figure 10.

7000 (a) (c) P1 6000 Pb O 5000

4000

3000 S Ca

Intensity(cps) 2000 P1 Zn Si 1000

0

-1000 -101234567 Binding energy(keV)

Figure 9. Cont.

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10000 (b) (d) P2 9000 Metals 2020, 10, 465 12 of 14 Metals 2020, 10, x FOR PEER REVIEW 8000 Pb 12 of 14 7000

6000 O 500010000 (b) (d) P2 40009000

Intensity(cps) 30008000 Pb S P 20007000 2 Ca Zn 10006000 O Si 50000

-10004000 -101234567 Intensity(cps) 3000 BindingS energy(keV) P 2000 2 Ca Zn 1000 Si 0 -1000 -101234567 Binding energy(keV) Figure 9. SEM images of leaching residue (a) and lead concentrate (b); EDS images of leaching residueFigure ( 9.c) andSEM lead images concentrate of leaching (d residue). (a) and lead concentrate (b); EDS images of leaching residue (c) and lead concentrate (d).

6500 Figure 9. SEM images of leaching1-PbSO residue (a) and lead10000 concentrate(b) (b); EDS images of leaching 6000 (a) 4 1 1-PbSO4 1 2-3PbO·PbSO ·H O 9000 5500residue (c) and lead concentrate (d). 4 2 2-3PbO·PbSO4·H2O 5000 3-SiO 2 8000 3-SiO2 4500 2 4-ZnFe2O4 7000 4000 35006500 6000 1-PbSO 10000 (b) 30006000 (a) 4 5000 1 1-PbSO4 1 25005500 3 2-3PbO·PbSO ·H O 9000 2-3PbO·PbSO ·H O 4 2 4000 2 4 2 Intensity (a.u.)Intensity

Intensity (a.u.) Intensity 20005000 3-SiO2 8000 3-SiO 3000 3 2 15004500 2 4-ZnFe2O4 7000 10004000 2000 4 6000 5003500 1000 03000 5000 0 2500 3 -500 4000 2 5 1015202530354045505560 (a.u.)Intensity 5 1015202530354045505560

Intensity (a.u.) Intensity 2000 3000 3 1500 2θ (°) 2θ (°) 1000 2000 4 500 Figure 10. XRD diffractogram of leaching residue1000 (a) and lead concentrate (b). 0 0 -500 5 1015202530354045505560 It can be seen from Figures9 and 10 that after leaching, the5 amount 1015202530354045505560 of lead sulfate phase increased 2θ (°) 2θ (°) significantly,Figure while the 10. amountsXRD diffractogram of zinc oxide of leaching and lead residue oxide (a phases) and lead decreased concentrate significantly. (b). At the same time, some impurities such as calcium, magnesium and aluminum react with sulfuric acid to

formIt correspondingcan be seen from salts Figures that can 9 either and report10 that to after the residue,leaching, e.g., the , amount or of remain lead sulfate in the liquid phase increasedphase. In addition,significantly, after thewhile co-leaching the amounts of sulfuric of acidzinc and oxide sulfurous and acid,lead zincoxide was phases transformed decreased from Figure 10. XRD diffractogram of leaching residue (a) and lead concentrate (b). significantly.zinc oxide and At zincthe same ferrite time, into ZnSOsome4 impuritiesinto a liquid such phase, as calcium, while lead magnesium was transformed and aluminum into anglesite react withand sulfuric 3PbO PbSO acid toH formO in thecorresponding residue [26]. Thesalts characteristics that can either of report ZnSO toas the a soluble residue, e.g., and PbSOgypsum,and or It can· be seen4· 2 from Figures 9 and 10 that after leaching, the4 amount of lead sulfate 4phase remain3PbO PbSOin the4 liquidH2O as phase. an insoluble In addition, salt were after utilized the co-leaching to realize of the sulfuric effective acid separation and sulfurous of lead andacid, zinc. zinc increased· significantly,· while the amounts of zinc oxide and lead oxide phases decreased was transformed from zinc oxide and zinc ferrite into ZnSO4 into a liquid phase, while lead was 4.significantly. Conclusions At the same time, some impurities such as calcium, magnesium and aluminum react transformedwith sulfuric into acid anglesite to form and corresponding 3PbO·PbSO salts4·H2O that in thecan residueeither report [26]. Theto the characteristics residue, e.g., ofgypsum, ZnSO4 asor a solubleremainBased saltin the onand liquid the PbSO results phase.4 and obtained 3PbO·PbSOIn addition, in this4 ·Hafter work,2O theas wean co-leaching drewinsoluble the following saltof sulfuric were utilized conclusions: acid and to sulfurousrealize the acid, effective zinc separation of lead and zinc. (1)was transformedAn oxidizing from roasting zinc cinder oxide whichand zinc contained ferrite into 16.9% ZnSO Pb, 30.5%4 into a Zn, liquid 10.3% phase, Fe and while 25.1% lead S was was transformed into anglesite and 3PbO·PbSO4·H2O in the residue [26]. The characteristics of ZnSO4 as 4. Conclusionsobtained from a Pb-Zn sulphide ore dressing plant in the Hanyuan area of China. Anglesite and a solublelead salt oxide and were PbSO the4 and main 3PbO·PbSO Pb-bearing4·H minerals.2O as an insoluble Zinc oxide, salt zinc were sulfate utilized and to zinc realize ferrite the were effective the separationBasedmain on Zn-bearingof the lead results and minerals. zinc. obtained in this work, we drew the following conclusions: (2) A sulfuric acid leaching—gravity shaking-table separation process was used to recover lead (1)4. ConclusionsAn oxidizing roasting cinder which contained 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S was obtainedand zinc from from a thePb-Zn oxidizing sulphide roasting ore dressing cinder. Theplant zinc in the leaching Hanyuan efficiency area of of China. 96.7% wasAnglesite obtained and leadBasedunder oxide on the werethe leaching results the main conditions obtained Pb-bearing in used: this minerals.awo leachingrk, we Zincdrew temperature oxide, the following zinc of 55sulfate◦ C;conclusions: a and leaching zinc timeferrite of were 90 min; the a sulfuric acid dosage of 20%; a sulfurous acid dosage of 4%; a cinder particle size of <0.3 mm; (1) mainAn oxidizingZn-bearing roasting minerals. cinder which contained 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S was and a solid-liquid ratio of R = 1:4. After the gravity shaking-table separation, a lead concentrate (2) A obtainedsulfuric acid from leaching—gravity a Pb-Zn sulphide shaking-tableore dressing plant separation in the Hanyuan process was area used of China. to recover Anglesite lead and with 50.2% Pb, 2.33% Zn and a lead recovery of 86.0% was produced. zinclead from oxide the were oxidizing the main roasting Pb-bearing cinder. minerals. The zinc leachingZinc oxide, efficiency zinc sulfate of 96.7% and waszinc obtainedferrite were under the main Zn-bearing minerals.

(2) A sulfuric acid leaching—gravity shaking-table separation process was used to recover lead and zinc from the oxidizing roasting cinder. The zinc leaching efficiency of 96.7% was obtained under

Metals 2020, 10, 465 13 of 14

(3) XRD, SEM and EDS analysis of the leaching residue and lead concentrate show that the main minerals phase in the leaching residue are anglesite, 3PbO PbSO H O, SiO and ZnFe O , · 4· 2 2 2 4 and the main minerals phase in lead concentrate are anglesite, 3PbO PbSO H O and SiO .A · 4· 2 2 small amount of gangue compounds such as SiO2 is generated in the lead concentrate due to mechanical entrainment.

Author Contributions: This is a joint work of the five authors; each author was in charge of their expertise and capability: J.X. for writing, formal analysis and original draft preparation, Y.P. and W.D. for conceptualization, T.C. for validation, K.Z. for methodology and J.X. for investigation. All authors have read and agreed to the published version of the manuscript. Funding: This research received no external funding. Acknowledgments: This work was supported by the Sichuan Science and Technology Program (Grant Nos. 2018FZ0092); Key Laboratory of Sichuan Province for Comprehensive Utilization of Vanadium and Titanium Resources Foundation (2018FTSZ35); and the China Geological Big Survey (Grant No. DD20190694). Conflicts of Interest: On behalf of all authors, the corresponding author states that there is no conflict of interest. The funders had no role in the design, analyses or interpretation of any data of the study.

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