A method for removing iron from zinc by zinc oxygen pressure leaching

Abstract

This invention relates to a method for removing iron from zinc through oxygen pressure leaching, comprising the following steps: performing a first-stage oxygen pressure leaching on the zinc concentrate to be treated, followed by solid-liquid separation to obtain a first-stage leaching residue and a first-stage leaching solution; performing a second-stage oxygen pressure leaching on the first-stage leaching residue, followed by solid-liquid separation to obtain a second-stage leaching residue and a second-stage leaching solution; returning the second-stage leaching solution to the first-stage oxygen pressure leaching process; performing hot acid leaching on the second-stage leaching residue, followed by solid-liquid separation to obtain hot acid residue and hot acid solution; neutralizing the hot acid solution to obtain neutralization residue and neutralization solution; returning the neutralization residue to the first-stage oxygen pressure leaching process; reducing the neutralization solution to obtain a reducing solution and reducing residue; performing hematite removal on the reducing solution, followed by solid-liquid separation to obtain iron-removed solution and iron slag; returning the reducing residue to the second-stage oxygen pressure leaching process; and returning the iron-removed solution to the first-stage oxygen pressure leaching process. This method can achieve open-circuit iron removal in the zinc-oxygen pressure leaching system while reducing the loss of valuable elements such as zinc, achieving a Zn recovery rate of over 99%.

Description

Technical Field

[0001] This invention relates to a method for removing iron from zinc by zinc oxygen pressure leaching, belonging to the field of hydrometallurgy. Background Technology

[0002] With the continuous depletion of high-zinc, low-iron zinc-containing minerals, zinc smelting enterprises are increasingly choosing zinc materials with lower zinc grades and higher iron content, such as high-iron zinc sulfide concentrate. This brings a series of new problems to existing zinc smelting enterprises: heavy iron handling load, high content of valuable metals carried away by the iron slag generation process, and low comprehensive recovery efficiency of copper and other valuable metals. Current pressure leaching technology for zinc concentrate focuses on improving zinc leaching efficiency. In the iron removal process, either in-reactor iron removal or external goethite iron removal is used. The main problem is the large volume of zinc-containing solution requiring iron removal, which leads to increased steam consumption, increased neutralizing agent consumption, and increased iron slag volume. Moreover, the low iron content in the resulting iron slag results in poor comprehensive iron recovery, and the overall efficiency of zinc smelting enterprises needs to be improved. Therefore, there is an urgent need to develop more advanced zinc oxygen pressure leaching technology that can achieve efficient zinc leaching, reduce the volume of iron removal, improve iron removal efficiency, achieve efficient leaching of other valuable metals such as copper, take into account the autothermal reaction of pressure leaching, and improve the economic benefits of zinc smelting systems.

[0003] Chinese invention patent application CN118345248A discloses a method for opening up iron from a high-iron sphalerite oxygen pressure leaching system. The method involves first finely grinding the high-iron sphalerite, then performing a first-stage oxygen pressure leaching to obtain a first-stage leaching slurry. This slurry is then thickened to obtain a thickened underflow and supernatant. The thickened underflow is then roughed to obtain a primary roughing concentrate and tailings. The primary roughing tailings are then scavenged to obtain a primary scavenging concentrate and tailings. The primary scavenging tailings are then subjected to a secondary scavenging to obtain a secondary scavenging concentrate and final iron-bearing tailings. The roughing concentrate, primary scavenging concentrate, and secondary scavenging concentrate are mixed and then subjected to a mixed cleaning process to obtain a final zinc-sulfur concentrate and primary cleaning tailings. The primary cleaning tailings are then returned to the roughing operation to form a closed loop. The first-stage oxygen pressure leaching slurry thickened supernatant enters the zinc smelting neutralization and iron removal process. While this technology solves the problem of open-circuit iron in existing iron removal methods that are not applicable to oxygen pressure leaching systems, it relies on a relatively complex flotation process and requires strict control of flotation concentration and flotation pulp temperature to enhance the separation of sulfur, zinc, and iron, which further increases the operational difficulty. In addition, the iron-bearing tailings obtained ultimately contain more than 30% iron, more than 2.5% zinc, and more than 17% sulfur, with a low iron grade and high zinc and sulfur loss rates, resulting in low comprehensive utilization value. Summary of the Invention

[0004] To address the shortcomings of existing technologies, the present invention aims to provide a method for removing iron from zinc using zinc-oxygen pressure leaching, thereby reducing the loss of valuable elements such as zinc while removing iron from the open-circuit zinc-oxygen pressure leaching system.

[0005] To solve the above-mentioned technical problems, the technical solution of the present invention is as follows:

[0006] A method for removing iron by zinc oxygen pressure leaching includes the following steps:

[0007] S1. After a first-stage oxygen pressure leaching of the zinc concentrate to be treated, solid-liquid separation is performed to obtain a first-stage leaching residue and a first-stage leaching solution.

[0008] The zinc concentrate is zinc sulfide concentrate;

[0009] S2. After performing two-stage oxygen pressure leaching on the first leaching residue, solid-liquid separation is performed to obtain a second leaching residue and a second leaching solution.

[0010] S3. Return the second leaching solution to the oxygen pressure leaching process in S1.

[0011] After hot acid leaching of the two leaching residues, solid-liquid separation is performed to obtain hot acid residue and Fe-rich residue. 3+ Hot acid solution;

[0012] S4. After neutralizing the hot acid solution, the solid and liquid are separated to obtain a neutralized residue and a neutralized solution with a pH value of 2 to 4.

[0013] S5. Return the neutralized residue to the oxygen pressure leaching process in S1;

[0014] After reducing the neutralized liquid with a reducing agent, solid-liquid separation is performed to obtain a reduced liquid and a reduced residue.

[0015] The reducing agent is ZnS and / or zinc concentrate;

[0016] S6. After removing iron from the reducing solution using the hematite method, solid-liquid separation is performed to obtain iron-removed liquid and iron slag.

[0017] The reduction residue is returned to the second-stage oxygen pressure leaching process in S2;

[0018] S7. The iron removal liquid is returned to the oxygen pressure leaching process in S1.

[0019] Therefore, zinc concentrate is mixed with returned secondary leaching solution, neutralization slag, iron removal solution, etc., and after a first-stage oxygen pressure leaching process, primary leaching residue and primary leaching solution are obtained. The primary leaching solution can be sent downstream for pre-neutralization, purification, electrolysis, casting, and other processes. The primary leaching residue is further mixed with reducing slag, first acid solution, etc., and after a second-stage oxygen pressure leaching process, secondary leaching solution and secondary leaching residue are obtained. Subsequently, valuable elements such as iron and zinc in the secondary leaching residue are fully leached out through hot acid leaching. The hot acid solution is then neutralized to prepare for the next reduction process, preventing the generation of toxic gases such as H2S during the reduction process using ZnS and / or zinc concentrate as reducing agents, and ensuring that sulfur exists in the reduction slag in the form of elemental S. After reduction, the reducing solution is subjected to hematite removal to obtain high-grade iron slag and iron removal liquid. The iron slag can be sold or further processed for resource utilization, thus opening the iron in the zinc-oxygen pressure leaching system. The iron removal liquid is returned to the first-stage oxygen pressure leaching process in S1, where the acid can participate in the first-stage oxygen pressure leaching. Valuable metals such as zinc can be further enriched in the first-stage leaching solution and enter subsequent processes, effectively reducing the loss of valuable metals such as zinc. At the same time, the reduction slag is returned to the second-stage oxygen pressure leaching process in S2. The reduction slag contains elemental S and ZnS, which has high activity after reduction but has not participated in the reaction. Due to the increase in specific surface area and activity, the reactions (1) and (6) in the second-stage oxygen pressure leaching process are easier to carry out and release heat stably to maintain the second-stage oxygen pressure leaching reaction. This helps to reduce the external heating required for the second-stage oxygen pressure leaching process. For example, the steam consumption can be reduced from 0.1t steam / t-zinc concentrate to 0.01t steam / t-zinc concentrate.

[0020] Furthermore, during the oxygen pressure leaching process, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced, and the initial acid concentration (based on the concentration of H2SO4) of the reaction system is controlled at 30–60 g / L, preferably 35–55 g / L, the reaction temperature is 140–160 °C, preferably 145–155 °C, the total pressure is 1.0–1.3 MPa, preferably 1.1–1.2 MPa, and the reaction time is 1–2 h, preferably 1.2–1.8 h.

[0021] Preferably, during the oxygen pressure leaching process, a surfactant is added to the reaction system, wherein the amount of surfactant added is 3-5‰ of the zinc concentrate; preferably, the surfactant is lignin.

[0022] During the two-stage oxygen pressure leaching, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced, and a first solution containing sulfuric acid is added. The initial molar ratio of acid to zinc (molar ratio of sulfuric acid to zinc) in the reaction system is controlled to be 1.6–2.6:1, preferably 1.7–2.5:1, the reaction temperature is 140–160℃, preferably 145–155℃, the total pressure is 1.0–1.3 MPa, preferably 1.1–1.2 MPa, and the reaction time is 1.5–3 h, preferably 1.8–2.8 h.

[0023] Preferably, during the two-stage oxygen pressure leaching, a surfactant is added to the reaction system, wherein the amount of surfactant added is 3-5‰ of the zinc concentrate corresponding to the first leaching residue; preferably, the surfactant is lignin.

[0024] Preferably, the final acid concentration of the second immersion solution is 35–65 g / L;

[0025] Preferably, the sulfuric acid concentration in the first solution is 140-180 g / L. More preferably, the first solution is waste electrolyte, thereby effectively disposing of waste electrolyte from the zinc smelting process and avoiding the introduction of other impurities.

[0026] Furthermore, in S3, during hot acid leaching, the leaching temperature is controlled at 80–99°C, preferably 90–95°C;

[0027] Preferably, the second leaching residue is mixed with a second solution containing sulfuric acid and subjected to hot acid leaching; more preferably, the sulfuric acid concentration in the second solution is 140-180 g / L, and the solid-liquid ratio of the second leaching residue to the second solution is 1 g: 2-5 mL; even more preferably, the second solution is waste electrolyte.

[0028] Preferably, the amount of the second solution added in S3 is 35-45 vol% of the total amount of the first solution in S2 and the second solution in S3, more preferably 38-42 vol%. This allows for the simultaneous satisfaction of oxygen pressure leaching and hot acid leaching while significantly reducing the volume of the solution requiring iron removal, lowering the heat consumption for iron removal, and consequently significantly reducing the cost of iron removal in the iron source opening process.

[0029] Furthermore, if the Cu content in the hot acid slag is less than 0.2%, it can be further subjected to flotation to recover valuable elements such as Cu.

[0030] Furthermore, in S4, zinc oxide and / or zinc calcined sand are used to neutralize the hot acid solution; thus, neutralization can be achieved while avoiding the introduction of other elements.

[0031] Preferably, the reaction temperature is controlled at 80-90°C during the neutralization process;

[0032] Preferably, the reaction time is controlled to be 60-90 minutes during the neutralization process.

[0033] Furthermore, in S5, during the reduction process, the amount of reducing agent added is such that the Fe in the neutralized solution... 3+ Completely converted to Fe 2+ The required amount of reducing agent is 1.2-1.8 times, preferably 1.35-1.65 times;

[0034] Preferably, during the reduction treatment, the reaction temperature is controlled at 70–80°C and the reaction time is 4–6 hours.

[0035] Furthermore, in S6, during the iron removal process using hematite, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced into the reducing solution, and the reaction temperature is controlled at 180-200℃, preferably 185-195℃, and the total pressure is 1.6-2.0 MPa, preferably 1.7-1.9 MPa; preferably, the reaction time is 4-6 h, more preferably 4.5-5.5 h.

[0036] The zinc concentrate contains 40-60% Zn, 4-12% Fe, and 20-40% S; preferably, the zinc concentrate contains 48-51% Zn, 6-9% Fe, 28-32% S, and 0.4-0.7% Cu.

[0037] The main chemical reactions that may be involved in this invention are as follows:

[0038] 2ZnS+O2+2H2SO4=2ZnSO4+2H2O+2S (1)

[0039] CuFeS2+O2+2H2SO4=CuSO4+FeSO4+2S+2H2O (2)

[0040] 4FeSO4+O2+2H2SO4=2Fe2(SO4)3+H2O (3)

[0041] ZnO + H₂SO₄ = ZnSO₄ + H₂O (4)

[0042] ZnS+Fe2(SO4)3=ZnSO4+S+2FeSO4 (5)

[0043] S + H₂O + O₂ = H₂SO₄ (6)

[0044] 4FeSO4+O2+4H2O=2Fe2O3+4H2SO4 (7)

[0045] 3Fe2(SO4)3+14H2O=2H3OFe3(SO4)2(OH)6+5H2SO4 (8)

[0046] 2H3OFe3(SO4)2(OH) 6+ H₂SO₄=3Fe₂(SO₄)₃+14H₂O (9)

[0047] Furthermore, the zinc concentrate contains more than 97 wt% of solids with a particle size of -44 μm.

[0048] Furthermore, the oxygen concentration of the oxygen-enriched gas is 80-99 vol%, and even further, 98-99 vol%.

[0049] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0050] (1) The method of the present invention can reduce the loss of valuable elements such as zinc while realizing the iron open circuit in the zinc oxygen pressure leaching system, and the Zn recovery rate can reach more than 98.5%.

[0051] (2) This invention uses hot acid leaching to remove iron concentrated in the secondary leaching slag. The volume of hot acid produced by hot acid leaching is relatively small, which greatly reduces the volume of subsequent iron removal solution. The heat required for the subsequent hematite iron removal process is significantly reduced, thereby significantly reducing the cost of iron removal. The high-temperature, high-pressure, and efficient iron removal of the iron-containing solution simultaneously achieves the technical effect of iron resource utilization. The iron grade in the resulting iron slag is higher than 56%, the impurity content is low, and the iron resource utilization rate is high, reaching over 75%. Moreover, the zinc content in the iron slag is ≤0.2%, and the sulfur content is ≤1.2%, resulting in minimal loss of valuable elements such as zinc and sulfur in the zinc smelting system.

[0052] (3) The present invention uses zinc sulfide and / or zinc concentrate to reduce the neutralizing solution, and the resulting reduction residue is returned to the second-stage oxygen pressure leaching process, which makes the second-stage oxygen pressure leaching reaction easier to carry out and helps to reduce the consumption of external heating, that is, saves the amount of steam used to provide external heating in the oxygen pressure leaching process. Attached Figure Description

[0053] Figure 1 This is a flowchart of a zinc-oxygen pressure leaching process for iron removal according to the present invention. Detailed Implementation

[0054] The present invention will be described in detail below with reference to embodiments. It should be noted that, unless otherwise specified, the embodiments and features described in the embodiments of the present invention can be combined with each other. Unless otherwise specified, the relevant percentages refer to mass percentages.

[0055] Example 1

[0056] The zinc-oxygen pressure immersion zinc iron removal method of this embodiment includes the following steps:

[0057] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 Mix 400g of lignin (1.6g), neutralization residue (16.33g, Zn 53%, Fe 9%, S 1.5%), add lignin (1.6g), iron-removed liquid (1L, H2SO4 40g / L, Zn 72.6g / L, Fe 1.5g / L, Cu 0.25g / L), and second leaching liquid (1.4L, H2SO4 42g / L, Zn 99.95g / L, Fe 10.28g / L, Cu 0.25g / L). The sample was subjected to oxygen-enriched gas with an oxygen concentration of 99 vol% (0.65 g / L) for a first-stage oxygen pressure leaching process. The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization and liquid-solid separation, a leaching residue (275 g, Zn 10.85%, Fe 22.65%, Cu 0.3%, S 40.88%) and a leaching solution (2.4 L, H₂SO₄ 12.5 g / L, Zn 161.4 g / L, Fe (total iron) 1.6 g / L, Fe 2+ 0.52 g / L, Cu 1.08 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0058] (2) Take all the above-mentioned leaching residue and reducing residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m²) 2 / cm 3 Mix the ingredients, add lignin (1.6g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the acid-zinc molar ratio at 1.78. Introduce oxygen-enriched gas (99 vol%) for a two-stage oxygen pressure leaching process. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, with a reaction time of 2 hours, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (290.7g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching solution (1.4L, H2SO4 42g / L, Zn 99.95g / L, Fe 10.28g / L, Fe...). 2+ 1.1 g / L, Cu 0.65 g / L), of which the steam consumption during the two-stage oxygen pressure leaching process is 0.01 t steam / t zinc concentrate.

[0059] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0060] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (290.7 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 45.33 g / L, Fe 50.09 g / L, Fe...). 2+ 0.8 g / L, Cu 0.17 g / L) and hot acid slag (116.3 g, Zn 0.5%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0061] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.33g / L, Fe 50.09g / L, Fe 2+ 32.65 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, a neutralization residue (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 63.01 g / L, Fe 51.09 g / L, Cu 0.17 g / L) were obtained. 2+ 0.7 g / L, Cu 0.26 g / L).

[0062] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0063] The above neutralization solution (1L, Zn 63.01g / L, Fe 51.09g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.26 g / L) was used for reduction treatment. 140 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) and a reducing solution (1 L, Zn 72.60 g / L, Fe 53.15 g / L, Fe...). 2+ 53.02 g / L, Cu 0.25 g / L).

[0064] (6) The above-mentioned reduction residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the second stage oxygen pressure leaching process;

[0065] The reduced solution (1L, Zn 72.61g / L, Fe 53.15g / L, Fe) was then subjected to further treatment. 2+53.02 g / L, Cu 0.26 g / L) were added to the high-pressure reactor, and oxygen with a purity of 99 vol% was introduced at 190℃. The total pressure was controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain iron removal liquid (1 L, Zn 72.60 g / L, Fe 1.5 g / L, Cu 0.25 g / L, H2SO4 40 g / L) and iron slag (90.93 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. Among them, the steam consumption of the iron removal process was 1.0 t steam / t zinc concentrate. (7) The above iron removal liquid was returned to the first oxygen pressure leaching process.

[0066] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 79.0%.

[0067] Comparative Example 1

[0068] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0069] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 400g of lignin (1.6g) was added, along with the iron-removed liquid (1L, H2SO4 70g / L, Zn 65.39g / L, Fe 4.5g / L, Cu 0.23g / L) and the second leaching liquid (1.4L, H2SO4 63g / L, Zn 110.04g / L, Fe 9.72g / L, Cu 0.76g / L). A first-stage oxygen pressure leaching was performed using 99 vol% oxygen gas at a reaction temperature of 150℃, a total pressure of 1.25MPa, and a reaction time of 1.5h. After cooling, depressurization, and liquid-solid separation, the following were obtained: first leaching residue (272g, Zn 9.85%, Fe 11.65%, Cu 0.3%, S 40.88%) and first leaching liquid (2.4L, H2SO4 46.8g / L, Zn 110.04g / L, Fe 9.72g / L, Cu 0.76g / L). 162.17 g / L, Fe 8.88 g / L, Fe 2+ 4.85 g / L, Cu 1.11 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0070] (2) Take all the above-mentioned leaching residue and reducing residue (180g, Zn 40%, Fe 7.6%, Cu 0.53%, S 30%, specific surface area 4.65m²) 2 / cm 3Mix the ingredients, add lignin (1.6g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the zinc-to-acid molar ratio at 1.51. Introduce oxygen gas with a concentration of 99 vol% for a two-stage oxygen pressure leaching process. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, and the reaction time at 2h, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (352.56g, Zn 1.2%, Fe 10.01%, Cu 0.1%, S 51.61%) and the second leaching solution (1.4L, H2SO4 63g / L, Zn 110.04g / L, Fe 9.72g / L, Fe...). 2+ 0.9 g / L, Cu 0.76 g / L), the steam consumption for the two-stage oxygen pressure leaching is 0.015 t steam / t zinc concentrate.

[0071] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0072] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (352.56 g, Zn 1.2%, Fe 10.01%, Cu 0.1%, S 51.61%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 46.38 g / L, Fe 31.76 g / L, Fe...). 2+ 0.8 g / L, Cu 0.21 g / L) and hot acid slag (141.02 g, Zn 0.5%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0073] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 46.38g / L, Fe 31.76g / L, Fe 2+ The zinc concentrate (0.8 g / L, Cu 0.21 g / L) was reduced using zinc concentrate. 200 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (180 g, Zn 40%, Fe 7.6%, Cu 0.53%, S 30%) and a reducing solution (1 L, Zn 65.98 g / L, Fe 32.40 g / L, Fe...). 2+ 32.30 g / L, Cu 0.23 g / L).

[0074] (5) The above-mentioned reduction residue (180g, Zn 40%, Fe 7.6%, Cu 0.53%, S 30%) is returned to the second stage oxygen pressure leaching process;

[0075] The reduced solution (1L, Zn 65.98g / L, Fe 32.40g / L, Fe) was then subjected to further treatment. 2+ 32.30 g / L Zn, 0.23 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 65.95 g / L, Fe 4.5 g / L, Cu 0.23 g / L, H₂SO₄ 70 g / L) and iron slag (59.29 g, Zn 0.2%, Fe 53.25%, S 1.2%). The iron slag was sent for subsequent resource recovery. The iron removal process consumed 1.0 t of steam per t of zinc concentrate.

[0076] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0077] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 40.9%.

[0078] The results showed that skipping the pre-neutralization process and directly reducing the hot acid solution produced a strong rotten egg odor during reduction, leading to increased zinc concentrate usage and a deteriorated operating environment. Increasing the amount of zinc concentrate used for reduction also increased the initial acidity before iron removal from hematite, having little impact on the zinc leaching rate, but significantly increasing the iron content in the first-stage leaching solution, increasing the pressure on subsequent iron removal and purification. Furthermore, the overall iron resource recovery rate decreased significantly to 40.9%. Steam consumption also increased in the second-stage oxygen pressure leaching.

[0079] Comparative Example 2

[0080] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0081] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3Mix 400g of chelated iron, neutralization residue (15.51g, Zn 53%, Fe 9%, S 1.5%), add lignin (1.6g), iron-removed liquid (1L, H2SO4 40g / L, Zn 72.45g / L, Fe 1.5g / L, Cu 0.25g / L), and second leaching liquid (1.4L, H2SO4 42g / L, Zn 102.58g / L, Fe 10.38g / L, Cu...). The sample was subjected to oxygen leaching with an oxygen concentration of 99 vol% (0.67 g / L). The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization and liquid-solid separation, a leaching residue (275 g, Zn 10.85%, Fe 22.65%, Cu 0.3%, S 40.88%) and a leaching solution (2.4 L, H2SO4 12.5 g / L, Zn 162.68 g / L, Fe 1.6 g / L, Fe...) were obtained. 2+ 0.52 g / L, Cu 1.09 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0082] (2) Take all the above-mentioned leaching residue and reducing residue (138g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.66m²) 2 / cm 3 Mix the ingredients, add lignin (1.6g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the zinc-to-acid molar ratio at 1.70. Introduce oxygen gas with a concentration of 99 vol% for a two-stage oxygen pressure leaching. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, and the reaction time at 2h, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (297.36g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching solution (1.4L, H2SO4 42g / L, Zn 102.58g / L, Fe 10.38g / L, Fe...). 2+ 1.1 g / L, Cu 0.67 g / L), the steam consumption of the two-stage oxygen pressure leaching process is 0.011 t steam / t zinc concentrate.

[0083] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0084] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (297.36 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 45.38 g / L, Fe 51.23 g / L, Fe...).2+ 0.8 g / L, Cu 0.18 g / L) and hot acid slag (118.94 g, Zn 0.5%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0085] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.38g / L, Fe 51.23g / L, Fe 2+ 31.02 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 1. After liquid-solid separation, a neutralization residue (15.51 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 62.18 g / L, Fe 51.74 g / L, Cu 0.18 g / L) were obtained. 2+ 0.7 g / L, Cu 0.26 g / L).

[0086] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0087] The above neutralization solution (1L, Zn 62.18g / L, Fe 51.74g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.26 g / L) was reduced by adding 150 g of zinc concentrate at a reaction temperature of 85 °C for 300 min. After the reaction, the solid-liquid mixture was separated to obtain a reducing residue (138 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) and a reducing solution (1 L, Zn 72.47 g / L, Fe 53.15 g / L, Fe 0.26 g / L). 2+ 53.02 g / L, Cu 0.26 g / L).

[0088] (6) The above-mentioned reduction residue (138g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the second stage oxygen pressure leaching process;

[0089] The reduced solution (1L, Zn 72.47g / L, Fe 53.15g / L, Fe) was then subjected to further treatment. 2+53.02 g / L Zn, 0.26 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 72.45 g / L, Fe 1.5 g / L, Cu 0.25 g / L, H₂SO₄ 40 g / L) and iron slag (90.93 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption for the de-ironization process was 1.0 t steam / t zinc concentrate.

[0090] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0091] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 79.0%.

[0092] The results showed that lowering the pH control endpoint of the neutralization process resulted in the generation of a rotten egg smell during the reduction process, increased the amount of zinc concentrate used in the reduction process, had no significant effect on improving the zinc leaching rate and the iron resource utilization rate, increased the amount of reduction slag to be returned, and was not conducive to reducing the energy consumption of material transportation.

[0093] Comparative Example 3

[0094] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0095] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 Mix 400g of chelated iron, neutralization residue (19.59g, Zn 53%, Fe 12.5%, S 1.5%), add lignin (1.6g), iron-removed liquid (1L, H2SO4 40g / L, Zn 73.88g / L, Fe 1.5g / L, Cu 0.25g / L), and second leaching liquid (1.4L, H2SO4 42g / L, Zn 92.16g / L, Fe 10.01g / L, Cu...). 0.58 g / L), oxygen gas with an oxygen concentration of 99 vol% was introduced for a first-stage oxygen pressure leaching process. The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization and liquid-solid separation, a leaching residue (276 g, Zn 10.85%, Fe 22.65%, Cu 0.3%, S 40.88%) and a leaching solution (2.4 L, H2SO4 12.5 g / L, Zn 158.06 g / L, Fe 1.6 g / L, Fe 2+ 0.52 g / L, Cu 1.05 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0096] (2) Take all the above-mentioned leaching residue and reducing residue (101.2g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m²) 2 / cm 3 Mix the ingredients, add lignin (1.6g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the zinc-to-acid molar ratio at 2.06. Introduce oxygen gas with a concentration of 99 vol% for a two-stage oxygen pressure leaching process. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, and the reaction time at 2h, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (271.58g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching solution (1.4L, H2SO4 42g / L, Zn 92.16g / L, Fe 10.01g / L, Fe...). 2+ 1.1 g / L, Cu 0.58 g / L), of which the steam consumption of the two-stage oxygen pressure leaching process is 0.02 t steam / t zinc concentrate.

[0097] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0098] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (271.58 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 45.17 g / L, Fe 46.79 g / L, Fe...). 2+ 0.8 g / L, Cu 0.16 g / L) and hot acid slag (108.63 g, Zn 0.5%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0099] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.17g / L, Fe 46.79g / L, Fe 2+ 39.18 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 5. After liquid-solid separation, neutralization residue (19.59 g, Zn 53%, Fe 12.5%, Cu 0.57%, S 1.5%) and neutralization solution (1 L, Zn 66.39 g / L, Fe 42.95 g / L, Cu 0.16 g / L) were obtained. 2+ 0.7 g / L, Cu 0.26 g / L).

[0100] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0101] The above neutralization solution (1L, Zn 66.39g / L, Fe 42.95g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.26 g / L) was used for reduction treatment. 110 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, the reduced residue (101.20 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.58 m²) was obtained through liquid-solid separation. 2 / cm 3 ) and reducing solution (1L, Zn 73.94g / L, Fe 43.81g / L, Fe 2+ 43.75 g / L, Cu 0.26 g / L).

[0102] (6) The above-mentioned reduction residue (101.20g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the second stage oxygen pressure leaching process;

[0103] The reduced solution (1L, Zn 73.94g / L, Fe 43.81g / L, Fe) was then subjected to further treatment. 2+ 43.75 g / L Zn, 0.26 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the reactor was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 73.88 g / L, Fe 1.5 g / L, Cu 0.25 g / L, H₂SO₄ 40 g / L) and iron slag (74.95 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption for the de-ironization process was 1.0 t steam / t zinc concentrate.

[0104] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0105] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 71.1%.

[0106] The results showed that flocculent ferric hydroxide precipitate formed during the neutralization process in step (4), which caused the iron ions in the open circuit to the hot acid solution to disperse, reducing the iron resource utilization rate. In addition, the iron loss caused by the neutralization process further reduced the amount of concentrate used for subsequent zinc concentrate reduction, that is, reduced the amount of reduction slag produced, which in turn reduced the amount of reduction slag returned to the second-stage oxygen pressure leaching process, resulting in an increase in the amount of steam used in the second-stage oxygen pressure leaching process.

[0107] Comparative Example 4

[0108] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0109] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 400g of lignin (1.6g) was added, along with iron-removed liquid (1L, H2SO4 40g / L, Zn 79.99g / L, Fe 1.5g / L, Cu 0.25g / L) and second leaching liquid (1.4L, H2SO4 42g / L, Zn 99.81g / L, Fe 10.21g / L, Cu 0.64g / L). A first-stage oxygen pressure leaching was performed using 99 vol% oxygen gas at a reaction temperature of 150℃, a total pressure of 1.25MPa, and a reaction time of 1.5h. After cooling, depressurization, and liquid-solid separation, the following were obtained: first leaching residue (273g, Zn 10.85%, Fe 22.65%, Cu 0.3%, S 40.88%) and first leaching liquid (2.4L, H2SO4 12.5g / L, Zn 99.81g / L, Fe 10.21g / L, Cu 0.64g / L). 160.88 g / L, Fe 1.6 g / L, Fe 2+ 0.53 g / L, Cu 1.08 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0110] (2) Take all the above-mentioned leaching residue and reducing residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m²) 2 / cm 3 The following mixtures were prepared: neutralization residue (16.33g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%), lignin (1.6g), and 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). The acid-zinc molar ratio was 1.79. A two-stage oxygen pressure leaching was performed using 99 vol% oxygen gas. The reaction temperature was controlled at 150℃, the total pressure at 1.25 MPa, and the reaction time at 2 hours. After cooling, depressurization, and liquid-solid separation, the following were obtained: secondary leaching residue (289.30g, Zn 4.2%, Fe 21.01%, Cu 0.1%, S 51.62%) and secondary leaching solution (1.4L, H2SO4 42g / L, Zn 99.81g / L, Fe 10.21g / L, Fe...). 2+ 1.1 g / L, Cu 0.64 g / L), the steam consumption during the two-stage oxygen pressure leaching process is 0.011 t steam / t zinc concentrate.

[0111] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0112] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (289.30 g, Zn 4.2%, Fe 21.01%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 52.72 g / L, Fe 49.84 g / L, Fe...). 2+ 0.8 g / L, Cu 0.17 g / L) and hot acid slag (115.72 g, Zn 2.3%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0113] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 52.72g / L, Fe 49.84g / L, Fe 2+ 32.65 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, a neutralization residue (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 70.40 g / L, Fe 50.84 g / L, Cu 0.17 g / L) were obtained. 2+ 0.7 g / L, Cu 0.25 g / L).

[0114] (5) The neutralized residue is subjected to a second-stage oxygen pressure leaching;

[0115] The above neutralization solution (1L, Zn 70.40g / L, Fe 50.84g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.25 g / L) was used for reduction treatment. 140 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) and a reducing solution (1 L, Zn 80.01 g / L, Fe 51.25 g / L, Fe...). 2+ 51.15 g / L, Cu 0.26 g / L).

[0116] (6) The above-mentioned reduction residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the second stage oxygen pressure leaching process;

[0117] The reduced solution (1L, Zn 80.01g / L, Fe 51.25g / L, Fe) was then subjected to further treatment. 2+51.15 g / L Zn, 0.26 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the reactor was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 79.99 g / L, Fe 1.5 g / L, Cu 0.25 g / L, H₂SO₄ 40 g / L) and iron slag (87.68 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption for the de-ironization process was 1.0 t steam / t zinc concentrate.

[0118] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0119] Calculations show that throughout the entire process, the recovery rate of Zn was 98.9%, and the recovery rate of iron was 79.0%. The recovery rate of zinc decreased.

[0120] Comparative Example 5

[0121] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0122] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 400g of neutralization residue (16.33g, Zn 53%, Fe 9%, S 1.5%), and 128.8g of reducing residue (Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m²). 2 / cm 3 The mixture was stirred, and lignin (1.6 g) was added. Then, the iron-removed solution (1 L, H₂SO₄ 40 g / L, Zn 71.82 g / L, Fe 1.5 g / L, Cu 0.25 g / L) and the second leaching solution (1.4 L, H₂SO₄ 42 g / L, Zn 64.20 g / L, Fe 13.35 g / L, Cu 0.35 g / L) were added. A first-stage oxygen pressure leaching was performed by introducing oxygen gas with a concentration of 99 vol%. The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization, and liquid-solid separation, the first leaching residue (275 g, Zn 11.36%, Fe 22.65%, Cu 0.3%, S 40.88%) and the first leaching solution (2.4 L, H₂SO₄ 13.6 g / L, Zn 64.20 g / L, Fe 13.35 g / L, Cu 0.35 g / L) were added. 162.17 g / L, Fe 5.6 g / L, Fe 2+ 1.65 g / L, Cu 1.11 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0123] (2) All the above-mentioned first-leaching residues were mixed with lignin (1.6 g) and 1.4 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L). At this point, the acid-zinc molar ratio was 4.78. A second-stage oxygen pressure leaching was performed by introducing oxygen gas with an oxygen concentration of 99 vol%. The reaction temperature was controlled at 150 °C, the total pressure at 1.25 MPa, and the reaction time at 2 h. After cooling and depressurization and liquid-solid separation, the following were obtained: second-leaching residue (198 g, Zn 1.5%, Fe 16.08%, Cu 0.1%, S 51.62%) and second-leaching solution (1.4 L, H2SO4 73 g / L, Zn 64.2 g / L, Fe 21.36 g / L, Fe...). 2+ 2.1 g / L, Cu 0.35 g / L), the steam consumption of the second stage is 0.09 t steam / t zinc concentrate.

[0124] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0125] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (198 g, Zn 1.5%, Fe 16.08%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40.5 g / L, Zn 45.38 g / L, Fe 26.11 g / L, Fe 2+ 0.85 g / L, Cu 0.12 g / L) and hot acid slag (79.2 g, Zn 0.52%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0126] (4) Add the above hot acid solution (1L, H2SO4 40.5g / L, Zn 45.38g / L, Fe 26.11g / L, Fe 2+ 33.06 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, a neutralization residue (16.53 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 63.28 g / L, Fe 26.03 g / L, Cu 0.12 g / L) were obtained. 2+ 0.75 g / L, Cu 0.20 g / L).

[0127] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0128] The above neutralization solution (1L, Zn 63.28g / L, Fe 26.03g / L, Fe) was prepared using zinc concentrate. 2+The zinc concentrate (0.75 g / L, Cu 0.20 g / L) was used for reduction treatment. 140 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) and a reducing solution (1 L, Zn 72.88 g / L, Fe 27.16 g / L, Fe...). 2+ 27.02 g / L, Cu 0.20 g / L).

[0129] (6) The above-mentioned reduction residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the first oxygen pressure leaching process;

[0130] The reduced solution (1L, Zn 72.88g / L, Fe 27.16g / L, Fe) was then subjected to further treatment. 2+ 27.02 g / L Zn, 0.20 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the reactor was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 71.82 g / L, Fe 1.5 g / L, Cu 0.25 g / L, H₂SO₄ 40 g / L) and iron slag (46.47 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption for the de-ironization process was 1.0 t steam / t zinc concentrate.

[0131] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0132] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 57.31%.

[0133] Returning the reduction residue generated in step (5) to the first stage of oxygen pressure leaching has little effect on the zinc leaching rate, but it has a significant impact on the steam consumption of the second stage of oxygen pressure leaching process, which significantly increases the steam consumption and is not conducive to energy saving and consumption reduction.

[0134] Comparative Example 6

[0135] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0136] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3Mix 400g of lignin (1.6g), neutralization residue (32.65g, Zn 53%, Fe 9%, S 1.5%), add lignin (2L, H2SO4 30g / L, Zn 69.4g / L, Fe 1.5g / L, Cu 0.17g / L), and second leaching solution (0.4L, H2SO4 52g / L, Zn 179.24g / L, Fe 1.5g / L). Zn (36.20 g / L, Cu 2.27 g / L) was subjected to oxygen pressure leaching with 99 vol% oxygen gas at a reaction temperature of 150 °C, a total pressure of 1.25 MPa, and a reaction time of 1.5 h. After cooling, depressurization, and liquid-solid separation, a leaching residue (277 g, Zn 13.26%, Fe 22.65%, Cu 0.3%, S 40.88%) and a leaching solution (2.4 L, H₂SO₄ 13.4 g / L, Zn 132.36 g / L, Fe 1.6 g / L, Fe 2.27 g / L) were obtained. 2+ 0.53 g / L, Cu 1.15 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0137] (2) Take all the above-mentioned leaching residue and reducing residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m²) 2 / cm 3 The mixture was stirred, and lignin (1.6 g) was added. 0.4 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added. The acid-zinc molar ratio was controlled at 0.47. A two-stage oxygen pressure leaching was performed using oxygen gas with a concentration of 99 vol%. The reaction temperature was controlled at 150℃, the total pressure at 1.25 MPa, and the reaction time at 2 h. After cooling, depressurization, and liquid-solid separation, the following were obtained: leaching residue (292.18 g, Zn 4.5%, Fe 21.01%, Cu 0.1%, S 51.62%) and leaching solution (0.4 L, H2SO4 52 g / L, Zn 179.24 g / L, Fe 36.20 g / L, Fe...). 2+ 1.3 g / L, Cu 2.27 g / L), of which the steam consumption of the two-stage oxygen pressure leaching process is 0.01 t steam / t zinc concentrate.

[0138] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0139] 2 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (292.18 g, Zn 4.5%, Fe 21.01%, Cu 0.1%, S 51.62%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (2 L, H2SO4 40 g / L, Zn 46.94 g / L, Fe 25.17 g / L, Fe...).2+ 0.8 g / L, Cu 0.09 g / L) and hot acid slag (116.87 g, Zn 3.5%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0140] (4) Add the above hot acid solution (2L, H2SO4 40g / L, Zn 46.94g / L, Fe 25.17g / L, Fe 2+ 65.31 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, neutralization residue (32.65 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and neutralization solution (2 L, Zn 64.63 g / L, Fe 25.67 g / L, Cu 0.09 g / L) were obtained. 2+ 0.7 g / L, Cu 0.17 g / L).

[0141] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0142] The above neutralization solution (2L, Zn 64.63g / L, Fe 25.67g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.17 g / L) was used for reduction treatment. 140 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) and a reducing solution (2 L, Zn 69.43 g / L, Fe 26.19 g / L, Fe 0.17 g / L). 2+ 26.15 g / L, Cu 0.17 g / L).

[0143] (6) The above-mentioned reduction residue (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) is returned to the second stage oxygen pressure leaching process;

[0144] The reduced solution (2L, Zn 69.43g / L, Fe 26.19g / L, Fe) 2+26.15 g / L Zn, 0.17 g / L Cu were added to a high-pressure reactor. The reactor was heated to 190°C with 99 vol% oxygen introduced, and the total pressure controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain molten iron (2 L, Zn 69.4 g / L, Fe 1.5 g / L, Cu 0.17 g / L, H₂SO₄ 30 g / L) and iron slag (89.59 g, Zn 0.2%, Fe 55.65%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption for the iron removal process was 2.25 t steam / t zinc concentrate.

[0145] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0146] Calculations show that the recovery rate of Zn in the entire process was 98.4%, and the recovery rate of iron was 73.2%.

[0147] The results showed that controlling the volume of waste electrolyte in step (2) to 0.4L and the volume of waste electrolyte in the hot acid leaching process in step (3) to 2L resulted in a decrease in zinc recovery and a doubling of the volume of iron removal solution, significantly increasing the amount of steam used in the hematite iron removal process.

[0148] Comparative Example 7

[0149] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0150] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 Mix 400g of lignin (1.6g), neutralization residue (16.33g, Zn 53%, Fe 9%, S 1.5%), and add iron-removed liquid (1L, H2SO4 25g / L, Zn 63.7g / L, Fe 27.5g / L, Cu 0.21g / L), and second leaching liquid (1.4L, H2SO4 70g / L, Zn 65.22g / L, Fe 9.03g / L, Cu 0.21g / L). 0.39 g / L), oxygen gas with an oxygen concentration of 99 vol% was introduced for a first-stage oxygen pressure leaching process. The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization and liquid-solid separation, a leaching residue (272 g, Zn 10.92%, Fe 22.72%, Cu 0.3%, S 40.85%) and a leaching solution (2.4 L, H2SO4 12.7 g / L, Zn 137.48 g / L, Fe (total iron) 6.7 g / L, Fe 2+ 4.89 g / L, Cu 0.92 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0151] (2) Take all the above-mentioned leaching residue and reducing residue (18.40g, Zn 16.5%, Fe 7.5%, Cu 0.53%, S 45.3%, specific surface area 4.56m²) 2 / cm 3 Mix the ingredients, add lignin (1.2g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the acid-zinc molar ratio at 4.57. Introduce oxygen gas with a concentration of 99 vol% for a two-stage oxygen pressure leaching. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, and the reaction time at 2h, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (209.1g, Zn 1%, Fe 21.01%, Cu 0.1%, S 52.64%) and the second leaching solution (1.4L, H2SO4 70g / L, Zn 65.22g / L, Fe 9.03g / L, Fe...). 2+ 1.2 g / L, Cu 0.39 g / L), of which the steam consumption during the two-stage oxygen pressure leaching process is 0.06 t steam / t zinc concentrate.

[0152] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0153] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (209.1 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 52.64%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 44.67 g / L, Fe 36.02 g / L, Fe...). 2+ 0.9 g / L, Cu 0.13 g / L) and hot acid slag (83.64 g, Zn 0.5%, Fe 5.1%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0154] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 44.67g / L, Fe 36.02g / L, Fe 2+ 32.65 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, a neutralization residue (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 62.35 g / L, Fe 36.74 g / L, Cu 0.13 g / L) were obtained. 2+ 0.7 g / L, Cu 0.21 g / L).

[0155] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0156] The above neutralization solution (1L, Zn 62.35g / L, Fe 36.74g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.21 g / L) was used for reduction treatment. 20 g of zinc concentrate was added, the reaction temperature was 85℃, and the reaction time was 300 min. After the reaction, liquid-solid separation yielded a reducing residue (18.40 g, Zn 16.5%, Fe 7.5%, Cu 0.53%, S 45.3%) and a reducing solution (1 L, Zn 63.73 g / L, Fe 37.48 g / L, Fe...). 2+ 11.24 g / L, Cu 0.21 g / L).

[0157] (6) The above-mentioned reduction residue (18.40g, Zn 16.5%, Fe 7.5%, Cu 0.53%, S 45.3%) is returned to the second-stage oxygen pressure leaching process;

[0158] The reduced solution (1L, Zn 63.73g / L, Fe 37.48g / L, Fe) was then subjected to further treatment. 2+ 11.24 g / L Zn, 0.21 g / L Cu were added to a high-pressure reactor. Oxygen with a purity of 99 vol% was introduced at 190℃, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain a de-iron liquid (1 L, Zn 63.7 g / L, Fe 27.5 g / L, Cu 0.21 g / L, H₂SO₄ 25 g / L) and iron slag (17.85 g, Zn 0.2%, Fe 56.74%, S 1.2%). The iron slag was sent for subsequent resource recovery. The iron removal process consumed 2.1 t of steam per t of zinc concentrate.

[0159] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0160] Calculations show that the recovery rate of Zn in the entire process was 99.7%, and the recovery rate of iron was 20.8%.

[0161] The results showed that reducing the amount of zinc concentrate added in step (5) to 20g directly caused an increase in Fe in the reducing solution obtained in step (5). 2+ The content decreased significantly. In addition, since less S was generated in reaction (5), the reaction process of reaction (6) was limited after returning to the second stage oxygen pressure leaching, as well as the heat generated. Therefore, the amount of steam used in the second stage oxygen pressure leaching process increased significantly.

[0162] Comparative Example 8

[0163] The zinc-oxygen pressure leaching method for iron removal in this comparative example includes the following steps:

[0164] (1) The zinc concentrate (with a solid content greater than 97% of particle size -44μm, Zn 49%, Fe 8%, S 29%, Cu 0.5%, and a specific surface area of ​​3.75m²) was tested. 2 / cm 3 Mix 400g of lignin (1.6g), neutralization residue (16.33g, Zn 53%, Fe 9%, S 1.5%), and add iron-removed liquid (1L, H2SO4 40g / L, Zn 97.6g / L, Fe 1.6g / L, Cu 0.28g / L), and second leaching liquid (1.4L, H2SO4 39g / L, Zn 109.15g / L, Fe 10.92g / L, Cu 0.28g / L). 0.76 g / L), oxygen gas with an oxygen concentration of 99 vol% was introduced for a first-stage oxygen pressure leaching process. The reaction temperature was 150℃, the total pressure was 1.25 MPa, and the reaction time was 1.5 h. After cooling and depressurization and liquid-solid separation, a first leaching residue (275 g, Zn 10.85%, Fe 22.65%, Cu 0.3%, S 40.86%) and a first leaching solution (2.4 L, H2SO4 12.6 g / L, Zn 173.92 g / L, Fe (total iron) 1.7 g / L, Fe 2+ 0.5 g / L, Cu 1.16 g / L), the first immersion solution is sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0165] (2) Take all the above-mentioned leaching residue and reducing residue (184g, Zn 43%, Fe 7.7%, Cu 0.52%, S 30%, specific surface area 4.42m²) 2 / cm 3 Mix the ingredients, add lignin (1.6g), and add 1.4L of waste electrolyte (Zn 43g / L, H2SO4 160g / L). At this point, control the acid-zinc molar ratio at 1.37. Introduce oxygen gas with a concentration of 99 vol% for two-stage oxygen pressure leaching. Control the reaction temperature at 150℃ and the total pressure at 1.25MPa, and the reaction time at 2h, based on the steam flow rate. After cooling, depressurization, and liquid-solid separation, obtain the second leaching residue (330.48g, Zn 8.8%, Fe 21.01%, Cu 0.1%, S 51.63%) and the second leaching solution (1.4L, H2SO4 39g / L, Zn 109.15g / L, Fe 10.92g / L, Fe...). 2+ 0.9 g / L, Cu 0.76 g / L), of which the steam consumption during the second-stage oxygen pressure leaching process is 0.03 t steam / t zinc concentrate.

[0166] (3) Return all the second leaching solution to the first oxygen pressure leaching stage;

[0167] 1 L of waste electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added to the above-mentioned second-stage slag (330.48 g, Zn 8.8%, Fe 21.01%, Cu 0.1%, S 51.63%). After reacting at 92 °C for 2 h, solid-liquid separation was performed to obtain a hot acid solution (1 L, H2SO4 40 g / L, Zn 66.27 g / L, Fe 56.94 g / L, Fe...). 2+ 0.8 g / L, Cu 0.20 g / L) and hot acid slag (264.38 g, Zn 4.2%, Fe 5%, Cu 0.08%, S 90.2%), the hot acid slag is sent to the flotation process for further processing.

[0168] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 66.27g / L, Fe 56.94g / L, Fe 2+ 32.65 g of zinc calcinate (Zn 57%, Fe 8.5%, S 1%, Cu 0.55%) was slowly added to the solution at 85℃. The reaction was carried out for 75 min, and the final pH was controlled at 2.5. After liquid-solid separation, a neutralization residue (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S 1.5%) and a neutralization solution (1 L, Zn 83.95 g / L, Fe 58.07 g / L, Cu 0.20 g / L) were obtained. 2+ 0.7 g / L, Cu 0.28 g / L).

[0169] (5) The neutralized residue is subjected to a first-stage oxygen pressure leaching;

[0170] The above neutralization solution (1L, Zn 83.95g / L, Fe 58.07g / L, Fe) was prepared using zinc concentrate. 2+ The zinc concentrate (0.7 g / L, Cu 0.28 g / L) was reduced by adding 200 g of zinc concentrate. The reaction temperature was 85℃ and the reaction time was 300 min. After the reaction was completed, the reducing residue (184 g, Zn 43%, Fe 7.7%, Cu 0.52%, S 30%) and the reducing solution (1 L, Zn 89.83 g / L, Fe 59.24 g / L, Fe 0.28 g / L) were obtained by liquid-solid separation. 2+ 58.94 g / L, Cu 0.28 g / L).

[0171] (6) The above-mentioned reduction residue (184g, Zn 43%, Fe 7.7%, Cu 0.52%, S 30%) is returned to the second-stage oxygen pressure leaching process;

[0172] The reduced solution (1L, Zn 89.83g / L, Fe 59.24g / L, Fe) 2+58.94 g / L Zn, 0.28 g / L Cu were added to a high-pressure reactor. Oxygen with a purity of 99 vol% was introduced at 190℃, and the total pressure was controlled at 1.8 MPa. After reacting for 300 min, the mixture was cooled and depressurized to obtain a de-ironized liquid (1 L, Zn 89.78 g / L, Fe 1.6 g / L, Cu 0.28 g / L, H₂SO₄ 40 g / L) and iron slag (101.34 g, Zn 0.2%, Fe 56.75%, S 1.2%). The iron slag was sent for subsequent resource recovery. The steam consumption in the iron removal process was 1.05 t steam / t zinc concentrate.

[0173] (7) The above-mentioned iron removal liquid is returned to the oxygen pressure leaching process.

[0174] Calculations show that the recovery rate of Zn in the entire process was 96.1%, and the recovery rate of iron was 66%.

[0175] The results showed that increasing the amount of zinc concentrate added in step (5) to 200g, with the addition of excessive zinc concentrate, could satisfy the requirement for Fe. 2+ The reducing solution had a high content of zinc concentrate, but due to the excessive amount of unreacted zinc concentrate remaining, the overall specific surface area of ​​the reducing slag decreased, affecting the heat supply per unit time of the reaction (6). The steam consumption in the second-stage oxygen pressure leaching process increased to some extent. In addition, due to the excessive amount of unreacted zinc sulfide, the zinc content in the second-stage leaching slag increased, resulting in a decrease in the total zinc recovery rate. The increase in the total amount of zinc concentrate used led to a decrease in the iron resource utilization rate.

[0176] The above embodiments should be understood as being used only to illustrate the present invention more clearly, and not to limit the scope of the present invention. After reading the present invention, any modifications of the present invention in various equivalent forms by those skilled in the art fall within the scope defined by the appended claims.

Claims

1. A method for removing iron from zinc using zinc-oxygen pressure leaching, characterized in that, Includes the following steps: S1. After a first-stage oxygen pressure leaching of the zinc concentrate to be treated, solid-liquid separation is performed to obtain a first-stage leaching residue and a first-stage leaching solution. The zinc concentrate is zinc sulfide concentrate; during the first-stage oxygen pressure leaching, the initial acid concentration of the reaction system is controlled to be 30~60g / L; S2. After performing two-stage oxygen pressure leaching on the first leaching residue, solid-liquid separation is performed to obtain a second leaching residue and a second leaching solution. During the two-stage oxygen pressure leaching, a first solution containing sulfuric acid is added, and the initial molar ratio of sulfuric acid to zinc in the reaction system is controlled to be 1.6~2.6:

1. S3. Return the second leaching solution to the oxygen pressure leaching process in S1. After hot acid leaching of the two leaching residues, solid-liquid separation is performed to obtain hot acid residue and Fe-rich residue. 3+ Hot acid solution; S4. After neutralizing the hot acid solution, the solid and liquid are separated to obtain a neutralized residue and a neutralized solution with a pH value of 2 to 4. S5. Return the neutralized residue to the oxygen pressure leaching process in S1; After reducing the neutralized liquid with a reducing agent, solid-liquid separation is performed to obtain a reduced liquid and a reduced residue. The reducing agent is ZnS and / or zinc concentrate; the amount of reducing agent added is such that the Fe in the neutralizing solution is reduced. 3+ Completely converted to Fe 2+ The required amount of reducing agent is 1.2-1.8 times; S6. After removing iron from the reducing solution using the hematite method, solid-liquid separation is performed to obtain iron-removed liquid and iron slag. The reduction residue is returned to the second-stage oxygen pressure leaching process in S2; S7. The iron removal liquid is returned to the oxygen pressure leaching process in S1.

2. The method according to claim 1, characterized in that, During the oxygen pressure leaching process, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced, the reaction temperature is 140~160℃, the total pressure is 1.0~1.3 MPa, and the reaction time is 1~2 h.

3. The method according to claim 2, characterized in that, During a single oxygen pressure leaching process, a surfactant is added to the reaction system, wherein the amount of surfactant added is 3-5‰ of the zinc concentrate.

4. The method according to claim 3, characterized in that, The surfactant is lignin.

5. The method according to any one of claims 1-4, characterized in that, During the two-stage oxygen pressure leaching, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced, the reaction temperature is 140~160℃, the total pressure is 1.0~1.3 MPa, and the reaction time is 1.5~3 h.

6. The method according to claim 5, characterized in that, During the two-stage oxygen pressure leaching, a surfactant is added to the reaction system, wherein the amount of surfactant added is 3-5‰ of the zinc concentrate corresponding to the first leaching residue.

7. The method according to claim 6, characterized in that, The surfactant is lignin.

8. The method according to claim 5, characterized in that, The final acid concentration of the second immersion solution is 35~65 g / L.

9. The method according to claim 5, characterized in that, The sulfuric acid concentration in the first solution is 140-180 g / L.

10. The method according to claim 9, characterized in that, The first solution is waste electrolyte.

11. The method according to claim 5, characterized in that, In S3, during hot acid leaching, the leaching temperature is controlled at 80~99℃.

12. The method according to claim 11, characterized in that, In S3, during hot acid leaching, the leaching temperature is controlled at 90~95℃.

13. The method according to claim 11, characterized in that, The second leaching residue is mixed with a second solution containing sulfuric acid and subjected to hot acid leaching. The sulfuric acid concentration in the second solution is 140-180 g / L, and the solid-liquid ratio of the second leaching residue to the second solution is 1 g: 2-5 mL.

14. The method according to claim 13, characterized in that, The second solution is waste electrolyte.

15. The method according to claim 13, characterized in that, The amount of the second solution added in S3 is 35-45 vol of the total amount of the first solution in S2 and the second solution in S3.

16. The method according to claim 1, characterized in that, In S4, zinc oxide and / or zinc calcinate are used to neutralize the hot acid solution.

17. The method according to claim 16, characterized in that, During the neutralization process, the reaction temperature should be controlled at 80-90℃.

18. The method according to claim 16, characterized in that, When performing neutralization, the reaction time should be controlled to be 60-90 minutes.

19. The method according to claim 1, characterized in that, In S5, during the reduction process, the amount of reducing agent added is such that the Fe in the neutralized solution is reduced. 3+ Completely converted to Fe 2+ The required amount of reducing agent is 1.35-1.65 times.

20. The method according to claim 19, characterized in that, During the reduction process, the reaction temperature is controlled at 70~80℃ and the reaction time is 4~6h.

21. The method according to claim 1, characterized in that, In S6, when removing iron using the hematite method, oxygen-enriched gas with an oxygen concentration ≥50 vol% is introduced into the reducing solution, and the reaction temperature is controlled at 180-200℃, the total pressure at 1.6-2.0 MPa, and the reaction time at 4-6 h.

22. The method according to any one of claims 1-4 and 6-21, characterized in that, The zinc concentrate contains 40-60% Zn, 4-12% Fe, and 20-40% S.

23. The method according to claim 22, characterized in that, The zinc concentrate contains 48-51% Zn, 6-9% Fe, 28-32% S, and 0.4-0.7% Cu.

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