Extraction of Lead and Zinc from a Rotary Kiln Oxidizing Roasting Cinder

Abstract

In this study, sulfuric acid 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 ore in the Hanyuan area of China by a flotation-rotary kiln oxidizing roasting process. Anglesite and lead oxide were the main Pb-bearing minerals, while zinc sulfate, zinc oxide 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·PbSO₄·H₂O-dominated new lead mineral phase after acid leaching. A zinc 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 sulfuric acid 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·PbSO₄·H₂O, SiO₂ and ZnFe₂O₄, while the main chemical compounds in lead concentrate are anglesite, 3PbO·PbSO₄·H₂O and SiO₂.

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 metal recovery. Sulfide ore accounts for the majority of lead-zinc ore in China, as well 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 been derived, such as the direct selective flotation process, partial selective flotation process, floatable flotation process, asynchronous flotation process, partial bulk flotation process, fully bulk flotation process, branch series flotation process, gravity separation-floatation process and floatation-gravity separation process, etc. Sometimes a combined process of beneficiation and metallurgy is used to process extremely refractory bulk lead-zinc ores or lead-zinc oxide ores [1,2,3].

There are a large amount of Cu-Pb-Zn polymetallic sulfide ores in the Hanyuan area of China, of which the main metal-containing minerals are chalcopyrite, galena, sphalerite, marmatite and pyrite. The main gangue minerals are dolomite, calcite, quartz, feldspar, mica, chlorite, montmorillonite, etc. Assemblages of galena, sphalerite and marmatite (Fe-sphalerite) are currently produced as solid solutions with a fine grain size, and occur in these ores. Copper concentrate with more than 20% Cu, lead concentrate with more than 62% Pb and mixed Pb-Zn sulfide concentrate with about 40% Pb + Zn were obtained by floatation in the local dressing plant. The elements associated with the sulfide minerals are sulfur, silver, cadmium, gallium, indium, arsenic and antimony. The separation of lead and zinc from the lead-zinc sulfide ore is difficult using just flotation, due to the close association of lead with zinc. With the development of the market, the price of lead-zinc mixed concentrate has tended to decrease, while a large number of these concentrates are unsalable, neither of which are conducive to the improvement of economic benefits. Therefore, it is crucial that the effective separation of lead and zinc mixed lead-zinc concentrate is studied [4,5,6,7,8,9].

At present, sulfating roasting-leaching and high-pressure acid leaching-extraction separation are the main metallurgical methods to process Pb-Zn bulk concentrate. Zinc sulfate is obtained by the leaching and formation of crystal line products, while the lead sulfate remains in the leaching residue. 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 achieve effective separation [10,11,12,13,14,15,16]. In the local mining enterprises, lead-zinc bulk concentrate is treated in a rotary kiln, and the schematic diagram is shown in Figure 1.

The roasting temperature and roasting atmosphere are the two key factors, and are the most difficult to control in the oxidizing roasting process of mixed lead-zinc concentrate. The amount of air in and out of the blower is used to control the roasting atmosphere. If the oxygen content is sufficient, the roasting products are mainly sulfates. If the oxygen content is insufficient, the roasting products are mainly oxides. It is difficult to accurately control the roasting temperature of a rotary kiln, which is generally controlled by temperature installed sensors. The rotary kiln (B-C) region represents the high temperature belt, while the rotary kiln (C-D) region represents the middle temperature zone 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 metal oxide and metal sulfate in the oxidized roasting cinder. The separation of lead and zinc from oxidized calcined slag 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 hematite (Fe₂O₃) or magnetite (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_(s) + 3O_2(g) → 2PbO_(s) + 2SO_2(g)

(1)

PbO_(s) + SO_2(g) + 0.5O_2(g) → PbSO_4(s)

(2)

2ZnS_(s) + 3O_2(g) → 2ZnO_(s) + 2SO_2(g)

(3)

ZnO_(s) + SO_2(g) + 1/2O_2(g) → ZnSO_4(s)

(4)

(Zn,Fe)S_(s) + 2O_2(g) → ZnO_(s) + PbO_(s) + SO_2(g)

(5)

4FeS_2(s) + 7O_2(g) → 2Fe₂O_3(s) + 4SO_2(g)

(6)

3FeS_2(s) + 8O_2(g) → Fe₃O_4(s) + 6SO_2(g)

(7)

ZnO_(s) + Fe₂O_3(s) → ZnFe₂O_4(s)

(8)

CaCO_3(s) → CaO_(s) + CO_3(g)

(9)

CaMg(CO₃)_2(s) → CaO_(s) + MgO_(s) + 2CO_2(g)

(10)

CaO_(s) + Fe₂O_3(s) → CaFe₂O_4(s)

(11)

MgO_(s) + Fe₂O_3(s) → MgFe₂O_4(s)

(12)

CaO_(s) + SiO_2(s) → CaO·SiO_2(s)

(13)

MgO_(s) + SiO_2(s) → MgO·SiO_2(s)

(14)

CaO_(s) + Al₂O_3(s) → CaO·Al₂O_3(s)

(15)

MgO_(s) + Al₂O_3(s) → MgO·Al₂O_3(s)

(16)

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 quartz (SiO₂), calcium ferrite (CaFe₂O₄), magnesium ferrite (MgFe₂O₄), calcium aluminate (Ca₂Al₂O₄), etc. The chemical composition analysis of the sample is shown in

Table 1. Main chemical composition of the oxidizing roasting cinder (wt %).

Pb	Zn	Fe	S	As	CaO	SiO2	Al2O3	MgO	Na2O
16.9	30.5	3.36	15.1	0.01	3.22	8.36	4.23	2.89	1.66

. The lead and zinc mineral phase analyses for the main elements are shown in

Table 2. Lead phase analysis of the oxidizing roasting cinder (wt %).

Lead in Lead Sulfate	Lead in Lead Oxide	Lead in Lead Sulfide	Lead in Lead Carbonate
10.3	6.52	0.02	0.06

and

Table 3. Zinc phase analysis of the oxidizing roasting cinder (wt %).

Zinc in Zinc Sulfate	Zinc in Zinc Oxide	Zinc in Lead Sulfide	Zinc in Zinc Carbonate
20.2	10.2	0.06	0.04

, respectively, while a size fraction is shown in

Table 4. Size fraction analysis of the oxidizing roasting cinder.

Fractions (mm)	Content (wt %)	Grade (wt %)		Distribution (wt %)
		Pb	Zn	Pb	Zn
−2.00~+1.00	25.7	15.4	28.7	23.4	24.1
−1.00~+0.50	25.1	17.8	31.2	26.5	25.8
−0.50~+0.25	16.6	18.3	27.9	18.0	15.2
−0.25~+0.15	11.8	15.9	30.6	11.0	11.8
−0.15~+0.10	15.9	16.9	34.4	16.0	18.0
−0.10~0	4.90	17.3	31.8	5.10	5.10
Totals	100	16.9	30.5	100	100

, and the X-ray diffraction (XRD, X Pert pro, Panaco, the Netherlands) analysis is shown in Figure 2.

2.2. Chemical Reagent and Equipment

The main chemical reagents used in this test are sulfuric acid (purity 98%) and sulfurous acid (purity 99.5%), both obtained by Chengdu Jinshan Chemical Reagent Co., Ltd. (Chengdu, China). The main equipment used for the experiment was a magnetic stirrer (1.5 dm³, Jilin Exploration Machinery Factory, 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).

2.3. Experimental Procedure

Sulfuric acid leaching and gravity shaking-table separation were used to separate lead and zinc from the oxidizing roasting cinder, and this process is shown in Figure 3. To conduct acid leaching, raw materials (a mass of 100 g for each test) are mixed with the sulfuric acid in a 5–30% proportion and placed into the stirrer. Water was added to adjust the solid:liquid ratio to the target value, and the sulfuric acid reacts with the oxidizing roasting cinder to generate insoluble PbSO₄ and soluble ZnSO₄. After solid-liquid separation, ZnSO₄ reports to the lixivium and PbSO₄ to the leaching residue. Zinc-bearing lixivium is used for the preparation of ZnSO₄·7H₂O. Sulfuric acid leaching residue (a mass of 500 g for each test) and water were mixed in a 30–40% portion and gravity shaking-table to recover lead. The contents of lead and zinc in the sulfuric 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 analyzed to calculate the lead recovery. The calculation formula for zinc leaching efficiency is shown in Equation (17), and lead recovery is shown in Equation (18).

Zinc leaching efficiency = (Q₁ × α_Zn – Q₂ × θ_Zn)/Q₁ × α_Zn

(17)

Lead recovery = Q₃ × β_Pb/(Q₃ × β_Pb + Q₄ × θ_Pb)

(18)

where Q₁ is the weight of the oxidizing roasting cinder (g); Q₂ is the weight of the leaching residue (g); Q₃ is the weight of the gravity shaking-table separation concentrate (g); Q₄ is the weight of the gravity shaking-table separation tailings (g); α_Zn is content of the oxidizing roasting cinder (%); θ_Zn is the content of leaching residue (%); β_Pb is content of gravity shaking-table separation concentrate (%); and θ_Pb is content of gravity shaking-table separation tailings (%).

2.4. Analysis and Characterization

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

3. Results and Discussion

3.1. Separation of Zinc Using Sulfuric Acid Leaching

3.1.1. Sulfuric Acid Dosage

Hydrochloric acid, sulfuric acid and nitric acid are commonly used to extract lead and zinc by 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 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 separation. The effect of using different sulfuric acid dosages is shown in Figure 4a.

Results in Figure 4a confirm that increasing sulfuric acid dosage was conducive to improving zinc leaching efficiency and enriching 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 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 the leaching time is too long, it becomes easier to dissolve non-target elements in the sample. 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 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 that as the reaction time increases, the acid in the system is consumed, and there is little impetus for the reaction to continue. Therefore, a leaching time of 90 min was selected as most appropriate, the leaching efficiency of zinc was 84.8%, and the grades of lead and zinc in the leaching residue were 25.7% and 7.08%, respectively.

3.1.3. Leaching Temperature

The effects of different leaching temperatures are shown in Figure 5a. The results show that temperature has a significant influence on the leaching efficiency. When the leaching temperature rose to 55 °C, the leaching efficiency of zinc was 89.1%; when the leaching temperature was further increased to 75 °C, the increase of the leaching efficiency of zinc was relatively small. With the increase of temperature, the influence of sulfate ions in the solution on minerals increases, which is beneficial to the leaching of valuable metal zinc. When the temperature reached 55 °C, the leaching efficiency of other elements will increase significantly as the temperature continues to increase. Therefore, the leaching temperature was 55 °C, the leaching efficiency of zinc was 89.2%, the lead grade was increased to 28.8%, and the zinc grade was reduced to 5.66% in the leaching residue.

3.1.4. Fineness of the Oxidizing Roasting Cinder

The effects of different particle size fractions are shown in Figure 5b. Since the mixed lead-zinc concentrate was oxidized in a rotary kiln, zinc sulfate and lead sulfate are both present in the roasting cinder. If wet grinding is adopted to change the material size, it is not conducive to the control of the liquid-solid ratio in subsequent leaching tests and the concentration of Zn²⁺ in the leaching solution. Furthermore, it is not conducive to the subsequent separation or the recovery of zinc in the leaching solution. The results show that decreasing the particle size of the oxidizing roasting 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.

3.1.5. Solid-Liquid Ratio (R = m_solid:m_liquid)

The test results of different solid-liquid ratios are shown in Figure 6a. Once the solid-liquid ratio reached 1:4, the viscosity of the leaching solution changed little. Hence, diffusion rates of species to and from the reacting mineral surfaces did not increase significantly with the further decrease of the solid-liquid ratio. Hence, the leaching efficiency increased slowly. A solid-liquid ratio for the leaching of R = 1:4 was therefore suitable. The leaching efficiency of zinc was 91.4%, the lead grade was enriched to 32.8%, and the zinc grade was reduced to 5.08% in leaching residue.

3.1.6. Additive Dosage

The effects of different sulfurous acid dosage are shown in Figure 6b. The mineral analysis of the sample indicated the presence of ZnFe₂O₄, which does not readily react with sulfuric acid to produce ZnSO₄ and Fe₂(SO₄)₃ during the leaching process. Combining this with the previous relevant study, the addition of oxidizer in the leaching process was beneficial to improve the leaching efficiency. The predominance diagram of φ-pH for the SO₂-ZnFe₂O₄-H₂O system at different temperatures is shown in Figure 7. There are two mechanisms for the decomposition of ZnFe₂O₄ under acidic conditions. Under the oxidizing conditions, the first step of ZnFe₂O₄ decomposition is into Fe₂O₃ and Zn²⁺, and the second step is Fe³⁺ dissolution from Fe₂O₃. Under the reducing conditions, after ZnFe₂O₄ is decomposed into Fe₂O₃ and Zn²⁺, Fe₂O₃ is reductively dissolved to Fe²⁺. When the redox potential was controlled in the (B) region, zinc in ZnFe₂O₄ entered the solution in the form of Zn²⁺, while iron remained in the residue in the form of Fe₂O₃. When the redox potential was controlled in the (D), (E) and (F) regions, zinc and iron in ZnFe₂O₄ entered the solution in the form of Zn²⁺ and Fe²⁺, respectively [17,18,19,20].

The results in Figure 5b show that increasing the appropriate amount of sulfurous acid is conducive to improving the extraction efficiency of zinc. When the amount of sulfurous acid exceeded 4%, the leaching efficiency of zinc changes little. These results indicated that sulfurous acid can play 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.

3.1.7. Repeated Scale-up Test

In order to further validate test repeatability, repeated scale-up tests were carried out under the conditions employed as follows: a sulfuric acid dosage of 20%; a leaching time of 60 min; a leaching temperature of 55 °C; a cinder particle size of <0.3 mm; and a solid-liquid ratio of R = 1:3. These tests scaled to the ore mass up to 3000 g for each test, and the results are shown in

Table 5. Repeated scale-up test results (wt %).

Repeat	Contents in the Oxidizing Roasting Cinder		Contents in Leaching Residue		Leaching Efficiency of Zinc
	Pb	Zn	Pb	Zn
1	16.8	30.6	33.9	2.08	96.7
2	17.1	30.4	33.6	2.12	96.6
3	16.9	30.4	33.9	2.05	97.1
4	17.0	30.6	33.7	2.11	96.5
5	16.7	30.5	33.9	2.06	96.7
6	16.8	30.4	33.7	2.07	96.7
Average	16.9	30.5	33.8	2.08	96.7
Range ($R=E_{\max }-E_{\min }$) = 0.60
Arithmetic mean error ($\mathsf{\delta }=\frac{{\displaystyle \sum _{i=1}^N\left|d_i\right|}}{N}$) = 96.7
Sum square variation $SS={\displaystyle \sum _{i=1}^N{d}_i^2}={\displaystyle \sum _{i=1}^N{(E_i-\overline{E})}^2}$ = 0.21
Average deviation $MS=\frac{SS}{f}$ = 0.04
Standard deviation $s=\sqrt{\frac{{\displaystyle \sum {(E_i-\overline{E})}^2}}{N-1}}=\sqrt{MS}$ = 0.20

. These results show that the index of repeated experiments is superior to the conditions of results. Lead enters into the leaching residue, and is enriched in the leaching residue from 16.9% to 33.8%. The impurity elements also report to the residue at significant concentrations (

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

Pb	Zn	Fe	S	CaO	SiO2	Al2O3	MgO
33.8	2.07	0.26	5.12	9.18	36.9	0.25	0.45

).

3.2. Recovering Lead from Leaching Residue by Table Gravity Separation

Gravity shaking-table separation was adopted to further recover lead from the acid leaching residue, and the test procedure is shown in Figure 8. The test results are shown in

Table 7. Results of lead recovery lead from the leaching residues (wt %).

Products	Productivity	Grade		Recovery
		Pb	Zn	Pb	Zn
Lead concentrate	57.8	50.2	2.33	86.0	65.1
Tailings	42.2	11.2	1.71	14.0	34.9
Totals	100	33.8	2.07	100	100

, and the main element analysis results for the lead concentrate generated are shown in

Table 8. Main element analysis of lead concentrate (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

. The lead concentrate with a lead grade of 50.2% and recovery of 86.0% was obtained by two stages of grinding and three stages of gravity shaking-table separation. According to the quality standard for lead concentrate in China, the lead concentrate meets the requirement for level four lead concentrate, with a lead content of more than 50%.

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—mainly lead and zinc sulfate and oxides—exists in the form of different mineral phases, the oxidizing roasting cinder in different mineral dissolution mechanisms for the understanding of valuable metals such as lead, zinc and iron, magnesium, calcium and aluminum impurities such as leaching behavior have important significance. From the chemical composition, minerals composition and grain-sized analysis of the complex oxidizing roasting cinder, zinc was first extracted by sulphuric acid. The valuable metallic lead and zinc mainly exist in the mineral phases of lead sulfate, lead oxide, zinc oxide, zinc sulfate and zinc ferrite. At the same time, there are also corresponding chemical reactions for calcium, magnesium, aluminum compounds and sulfuric acid [21,22,23,24,25].

Therefore, it is very important to investigate the dissolution process of different mineral phases in acid solution. According to the X-ray diffraction (XRD) mineral analysis of the raw ore, the main mineral phases in the oxidizing roasting cinder were lead sulfate (PbSO₄), lead oxide (PbO), zinc sulfate (ZnSO₄), zinc oxide (ZnO), zinc ferrite (ZnFe₂O₄), calcium ferrite (CaFe₂O₄), magnesium ferrite (MgFe₂O₄) and calcium aluminate (Ca₂Al₂O₄). The main chemical reactions that may occur in the sulfuric acid leaching system were shown in Equations (19)–(24).

PbO_(s) + H₂SO_4(l) → PbSO_4(s) + H₂O_(l)

(19)

ZnO_(s) + H₂SO_4(l) → ZnSO_4(l) + H₂O_(l)

(20)

ZnFe₂O_4(s) + 4H₂SO_3(l) + 2O_2(g) → ZnSO_4(l) + Fe₂(SO₄)_3(l) + 4H₂O_(l)

(21)

CaFe₂O_4(s) + H₂SO_4(l) → CaSO_4(s) + 2FeSO_4(l) + H₂O_(l)

(22)

MgFe₂O_4(s) + 2H₂SO_4(l) → MgSO_4(l) + 2FeSO_4(l) + 2H₂O_(l)

(23)

Ca₂Al₂O_4(s) + 2H₂SO_4(l) → 2CaSO_4(s) + Al₂(SO₄)_3(l) +H₂O_(l)

(24)

XRD, scanning electron microscopy (SEM) and energy dispersive X-ray (EDS) are used to characterize the leaching residue and lead concentrate, and analyze the leaching mechanism of sulfuric acid system solution. SEM-EDS analysis results of leaching residue and lead concentrate are shown in Figure 9, and XRD analysis results for leaching residue and lead concentrate are shown in Figure 10.

It can be seen from Figure 9 and Figure 10 that after leaching, the amount of lead sulfate phase increased significantly, while the amounts of zinc oxide and lead oxide phases decreased significantly. At the same time, some impurities such as calcium, magnesium and aluminum react with sulfuric acid to form corresponding salts that can either report to the residue, e.g., gypsum, or remain in the liquid phase. In addition, after the co-leaching of sulfuric acid and sulfurous acid, zinc was transformed from zinc oxide and zinc ferrite into ZnSO₄ into a liquid phase, while lead was transformed into anglesite and 3PbO·PbSO₄·H₂O in the residue [26]. The characteristics of ZnSO₄ as a soluble salt and PbSO₄ and 3PbO·PbSO₄·H₂O as an insoluble salt were utilized to realize the effective separation of lead and zinc.

4. Conclusions

Based on the results obtained in this work, we drew the following conclusions:

(1)

An oxidizing roasting cinder which contained 16.9% Pb, 30.5% Zn, 10.3% Fe and 25.1% S was obtained from a Pb-Zn sulphide ore dressing plant in the Hanyuan area of China. Anglesite and lead oxide were the main Pb-bearing minerals. Zinc oxide, zinc sulfate and zinc ferrite were 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 the leaching conditions used: a leaching temperature of 55 °C; a leaching time of 90 min; a sulfuric acid 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 a lead recovery of 86.0% was produced.

(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·PbSO₄·H₂O, SiO₂ and ZnFe₂O₄, and the main minerals phase in lead concentrate are anglesite, 3PbO·PbSO₄·H₂O and SiO₂. A small amount of gangue compounds such as SiO₂ is generated in the lead concentrate due to mechanical entrainment.