Method for extracting iron by zinc immersion under zinc oxygen pressure

Abstract

The invention relates to a method for extracting iron by leaching zinc under oxygen pressure, which comprises the following steps: carrying out oxygen pressure leaching on zinc concentrate to be treated, and carrying out solid-liquid separation to obtain oxygen leaching residue and oxygen leaching liquid; after the oxygen leaching residues are subjected to hot acid leaching, solid-liquid separation is conducted, and hot acid residues and hot acid liquor rich in Fe < 3 + > are obtained; carrying out solid-liquid separation after neutralizing the hot acid liquid to obtain neutralized slag and neutralized liquid with the pH value of 2-4; returning the neutralized slag to the oxygen pressure leaching process in S1; carrying out reduction treatment on the neutralization solution by using a reducing agent, and carrying out solid-liquid separation to obtain a reduction solution and reduction slag; wherein the reducing agent is ZnS and / or zinc concentrate; after the reduction liquid is subjected to iron removal through a hematite method, solid-liquid separation is conducted, and iron removal liquid and iron slag are obtained; the reducing slag is returned to the S1 oxygen pressure leaching procedure; and returning the iron-removed liquid to the oxygen pressure leaching process in the step S1. By means of the method, loss of valuable elements such as zinc can be reduced while iron resource open circuit in the zinc oxygen pressure leaching system is achieved, and the Zn recovery rate can reach 98.9% or above.

Description

Technical Field

[0001] The invention relates to a method for extracting iron by zinc-oxygen pressure leaching, and belongs to the field of hydrometallurgy. Background Art

[0002] With the continuous reduction of high-zinc and low-iron zinc-containing minerals, zinc smelting enterprises have increasingly chosen zinc materials with reduced zinc grade and high iron content, such as high-iron sulfide zinc concentrate. This has brought a series of new problems to existing zinc smelting enterprises: heavy iron processing load, high content of valuable metals carried away by the iron slag production process, and low comprehensive recovery efficiency of other valuable metals such as copper. The current zinc concentrate pressure leaching process technology focuses on improving the leaching efficiency of zinc. In terms of the iron removal process, either iron precipitation in the kettle or iron removal with goethite outside the kettle is used. The main problem is that the volume of zinc-containing solution that needs to be iron-removed is huge, which leads to increased steam consumption, increased neutralizer consumption, and increased iron slag volume. In addition, the low iron content in the obtained iron slag results in poor comprehensive iron recovery effect, and the comprehensive benefits of zinc smelting enterprises need to be improved. The zinc smelting enterprises currently using zinc oxygen pressure leaching process mainly include one-stage oxygen pressure direct leaching and two-stage oxygen pressure staged leaching process. Compared with the two-stage oxygen pressure segmented leaching process, the main disadvantage of the one-stage oxygen pressure direct leaching process is that the sulfuric acid concentration and iron concentration in the oxygen pressure leaching solution are high, which brings great pressure to the subsequent acid reduction and iron removal. The existing technology usually considers the recovery or resource utilization of the iron in the leaching solution. Due to the large liquid-to-solid ratio and the large volume of the iron-containing solution, it will cause problems such as high energy consumption.

[0003] Therefore, there is an urgent need to develop more advanced zinc oxygen pressure leaching process technology, which can reduce the volume of liquid required to be processed for iron while achieving efficient zinc leaching, thereby improving the economic benefits of the zinc smelting system.

[0004] Chinese invention patent application specification CN118345248A discloses a method for extracting iron from a high-iron sphalerite oxygen pressure leaching system, which comprises the following steps: firstly, finely grinding the high-iron sphalerite, and then subjecting the high-iron sphalerite to a stage of oxygen pressure leaching to obtain a stage of leaching slurry, and then concentrating the stage of leaching slurry to obtain a concentrated underflow and a supernatant; then, the concentrated underflow is subjected to roughing to obtain a primary roughing concentrate and tailings, and the primary roughing tailings are subjected to a scavenging to obtain a primary scavenging concentrate and tailings, and the primary scavenging tailings are subjected to a secondary scavenging to obtain a secondary scavenging concentrate and a final iron-containing tailings; the roughing concentrate, the primary scavenging concentrate and the secondary scavenging concentrate are mixed and subjected to a mixed concentration to obtain a final zinc-sulfur concentrate and a primary concentrated tailings, and then the primary concentrated tailings are returned to the roughing operation to form a closed circuit, and the concentrated supernatant of the stage of oxygen pressure leaching slurry enters the zinc smelting and iron removal process. Although this technology solves the problem that the existing iron removal method is not suitable for open-circuit iron in the oxygen pressure leaching system, it relies on a more complex flotation process and requires strict control of the flotation concentration and the flotation slurry temperature to enhance the separation of sulfur, zinc and iron, which further increases the difficulty of operation; in addition, the iron-containing tailings finally obtained contain more than 30% iron, more than 2.5% zinc, and more than 17% sulfur, with a low iron grade, and a high zinc and sulfur loss rate, and a low comprehensive utilization value. Summary of the invention

[0005] In view of the shortcomings of the prior art, the purpose of the present invention is to provide a method for extracting zinc and iron by zinc-oxygen pressure leaching, which is based on a one-step oxygen pressure leaching method with a shorter process, so as to reduce the loss of valuable elements such as zinc while reclaiming the iron in the open-circuit zinc-oxygen pressure leaching system.

[0006] In order to solve the above technical problems, the technical solution of the present invention is as follows:

[0007] A method for extracting iron by zinc-oxygen pressure leaching, comprising the following steps:

[0008] S1. After oxygen pressure leaching of the zinc concentrate to be treated, solid-liquid separation is performed to obtain oxygen leaching slag and oxygen leaching liquid;

[0009] Wherein, the zinc concentrate is zinc sulfide concentrate;

[0010] S2, after hot acid leaching the oxygen leaching residue at 90-95°C, solid-liquid separation is performed to obtain hot acid slag and Fe-rich 3+ of hot acid;

[0011] S3, after neutralizing the hot acid liquid, separate the solid and the liquid to obtain a neutralized slag and a neutralized liquid with a pH value of 2 to 4;

[0012] S4, returning the neutralized slag to the oxygen pressure leaching process in S1;

[0013] After the neutralized liquid is reduced by a reducing agent, solid-liquid separation is performed to obtain a reduced liquid and reduced slag;

[0014] Wherein, the reducing agent is ZnS and / or zinc concentrate;

[0015] S5, after removing iron from the reducing liquid by the hematite method, separating the solid from the liquid to obtain iron removal liquid and iron slag;

[0016] Returning the reduced slag to the oxygen pressure leaching step S1;

[0017] S6, returning the iron removal solution to the oxygen pressure leaching process in S1.

[0018] Thus, the zinc concentrate is mixed with the returned neutralization slag, reduction slag, iron removal liquid and other materials, and after oxygen pressure leaching treatment, oxygen leaching slag and oxygen leaching liquid are obtained; the oxygen leaching liquid can be sent to the downstream side for pre-neutralization, purification, electrolysis, smelting and other processes, and the oxygen leaching slag is further treated by hot acid leaching. After the valuable elements such as iron and zinc in the oxygen leaching slag are fully leached, the hot acid liquid is neutralized to prepare for the next reduction treatment, so as to prevent the generation of toxic gases such as H2S during the reduction treatment process using ZnS and / or zinc concentrate as a reducing agent, and make the sulfur element exist in the reduction slag in the form of single substance S. After reduction, the reduction liquid is subjected to hematite method to remove iron, and high-grade iron slag and iron removal liquid can be obtained. The iron slag can be sold or further processed as a resource, realizing the resource recovery of iron in the zinc oxygen pressure leaching system; the iron removal liquid returns to the oxygen pressure leaching process in S1, and the acid therein can participate in the oxygen pressure leaching, and the valuable metals such as zinc therein can be further enriched in the oxygen leaching liquid and enter the subsequent process, effectively reducing the loss of valuable metals such as zinc. At the same time, the reduction slag returns to the oxygen pressure leaching process. The reduction slag contains single substance S and ZnS and other highly active substances that do not participate in the reaction after reduction. Due to the increase in its specific surface area, the activity increases, making it easier to carry out reactions (1) and (6) in the oxygen pressure leaching process, and stably releasing heat to maintain the progress of the oxygen pressure leaching reaction, which helps to reduce the external heat required for the oxygen pressure leaching process. For example, the steam consumption can be reduced from the conventional 800kg steam / t-zinc concentrate to less than 85kg steam / t-zinc concentrate.

[0019] Furthermore, in S1, the oxygen pressure leaching is a one-stage oxygen pressure leaching.

[0020] Furthermore, in S1, during oxygen pressure leaching, oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced, a first solution containing sulfuric acid is added, and the initial acid concentration of the reaction system is controlled to be 90-125 g / L, further 95-120 g / L, further 100-115 g / L, the reaction temperature is 135-158° C., further 140-155° C., further 145-150° C., the total pressure is 1.15-1.35 MPa, further 1.2-1.3 MPa, further 1.22-1.28 MPa, and the reaction time is 2-3.5 h, further 2.5-3 h;

[0021] Preferably, during oxygen pressure leaching, a surfactant is added to the reaction system, wherein the amount of the surfactant added is 3-6 wt‰ of the zinc concentrate, more preferably 4-5 wt‰; preferably, the surfactant is lignin;

[0022] Preferably, the concentration of sulfuric acid in the first solution is 140-180 g / L, preferably 145-175 g / L, and more preferably 150-170 g / L; more preferably, the first solution is waste electrolyte, thereby effectively absorbing waste electrolyte from the zinc smelting process and avoiding the introduction of other impurities.

[0023] Furthermore, in S2, when hot acid leaching is performed, the leaching temperature is controlled to be 91-94°C, preferably 92-93°C.

[0024] Furthermore, the oxygen leaching slag is mixed with a second solution containing sulfuric acid to carry out hot acid leaching; more preferably, the sulfuric acid concentration in the second solution is 140-180 g / L, preferably 145-175 g / L, and more preferably 150-170 g / L; the solid-liquid ratio of the oxygen leaching slag to the second solution is 1 g:2-6 mL, preferably 1 g:3-5 mL; more preferably, the second solution is waste electrolyte, thereby effectively absorbing the waste electrolyte of the zinc smelting process and avoiding the introduction of other impurities.

[0025] Preferably, the amount of the second solution added in S2 is 38-51 vol%, preferably 40-50 vol%, and more preferably 42-48 vol% of the total amount of the first solution in S1 and the second solution in S2. Thus, while satisfying the oxygen pressure leaching, valuable elements such as zinc in the oxygen leaching residue can be fully dissolved by hot acid leaching, and the volume of the solution to be deironed can be reduced, and the heat consumption required for deironing can be reduced, thereby significantly reducing the cost of deironing in the iron source opening process.

[0026] Furthermore, in S2, when hot acid leaching is performed, the initial acid concentration of the reaction system is 140-180 g / L, preferably 150-170 g / L, and preferably 155-165 g / L.

[0027] Furthermore, in S3, zinc oxide and / or zinc roasted sand is used to neutralize the hot acid solution, thereby achieving neutralization and avoiding the introduction of other elements.

[0028] Optionally, the hot acid slag can be subjected to further flotation.

[0029] Preferably, during the neutralization treatment, the reaction temperature is controlled to be 80-90°C;

[0030] Preferably, during the neutralization treatment, the reaction time is controlled to be 60-120 min.

[0031] Furthermore, in S4, when the reduction treatment is performed, the amount of the reducing agent added is such that Fe 3+ Completely converted to Fe 2+ 1.5-2.5 times, preferably 1.75-2.25 times, the amount of the required reducing agent;

[0032] Preferably, during the reduction treatment, the reaction temperature is controlled to be 70-80° C. and the reaction time is 4-6 hours.

[0033] Furthermore, in S5, when the hematite method is used for iron removal, an oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced into the reducing liquid, the reaction temperature is controlled to be 180-200°C, preferably 185-195°C, and the total pressure is 1.6-2.0 MPa, preferably 1.7-1.9 MPa; preferably, the reaction time is 4-6 h.

[0034] Furthermore, in the zinc concentrate, the content of Zn is 40-60%, the content of Fe is 4-12%, and the content of S is 20-40%;

[0035] Preferably, in the zinc concentrate, the content of Zn is 48-51%, the content of Fe is 6-9%, and the content of S is 28-32%. Optionally, the content of lead in the zinc concentrate is 1.5-5%.

[0036] The main chemical reaction equations that may be involved in the present invention are as follows:

[0037] 2ZnS+O2+2H2SO4=2ZnSO4+2H2O+2S (1)

[0038] CuFeS2+O2+2H2SO4=CuSO4+FeSO4+2S+2H2O (2)

[0039] 4FeSO4+O2+2H2SO4=2Fe2(SO4)3+H2O (3)

[0040] ZnO+H2SO4=ZnSO4+H2O (4)

[0041] ZnS+Fe2(SO4)3=ZnSO4+S+2FeSO4 (5)

[0042] S+H2O+O2=H2SO4 (6)

[0043] 4FeSO4+O2+4H2O=2Fe2O3+4H2SO4 (7)

[0044] 3Fe2(SO4)3+14H2O=2H3OFe3(SO4)2(OH)6+5H2SO4 (8)

[0045] 2H3OFe3(SO4)2(OH) 6+ H2SO4=3Fe2(SO4)3+14H2O (9)

[0046] Furthermore, in the zinc concentrate, the content of solid matter with a particle size of -37 μm is greater than 95 wt%.

[0047] Furthermore, the oxygen concentration of the oxygen-rich gas is 80-99 vol%, and further 98-99 vol%.

[0048] Compared with the prior art, the present invention has the following beneficial effects:

[0049] (1) The method of the present invention can realize the resource recovery of iron in the zinc oxygen pressure leaching system while reducing the loss of valuable elements such as zinc. The Zn recovery rate can reach more than 98.9%, and the process is shorter. Only one stage of oxygen pressure leaching is required to meet the high-efficiency recovery of valuable metals such as Zn.

[0050] (2) The present invention focuses on the iron and zinc in the oxygen leaching slag through hot acid leaching, ensuring the efficient leaching of zinc in the entire treatment process. Moreover, the volume of the hot acid solution generated by the hot acid leaching is small, which reduces the volume of the subsequent iron removal solution, and the heat required for the subsequent hematite method iron removal process is reduced, thereby reducing the iron removal cost. The iron-containing solution is efficiently iron-removed by high temperature and high pressure, and the technical effect of iron resource utilization is achieved. The iron grade in the obtained iron slag is higher than 56%, the impurity content is low, and the iron resource utilization rate is high, reaching more than 75%. The zinc content in the iron slag is ≤0.19%, and the S content is ≤1.2%. The loss of valuable elements such as zinc and sulfur in the zinc smelting system is small.

[0051] (3) The present invention uses zinc sulfide and / or zinc concentrate to reduce the neutralized solution, and the obtained reduced slag is returned to the oxygen pressure leaching process, which can make the oxygen pressure leaching reaction easier to carry out and help reduce the external heat consumption, that is, save the consumption of steam used to provide external heat in the oxygen pressure leaching process. BRIEF DESCRIPTION OF THE DRAWINGS

[0052] Figure 1 The present invention discloses a flow chart of a process for extracting iron by zinc-oxygen pressure leaching. DETAILED DESCRIPTION

[0053] The present invention will be described in detail below in conjunction with the embodiments. It should be noted that, in the absence of conflict, the embodiments and features in the embodiments of the present invention can be combined with each other. Unless otherwise specified, the relevant percentages refer to mass percentages.

[0054] Example 1

[0055] See also Figure 1 The method for removing iron by zinc-oxygen pressure leaching in this embodiment comprises the following steps:

[0056] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 24.49 g of neutralized slag (Zn 53%, Fe 9%, S1.5%, Pb 5.05%), and 105.6 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.72 m 2 / cm 3 ), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 81.16g / L Zn, 1.5g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 291.55g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.46g / L Zn, 5.6g / L Fe, 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 80kg steam / t zinc concentrate.

[0057] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0058] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0059] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 291.55 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.66 g / L, Fe28.69 g / L, Fe 2+ 0.7g / L) and hot acid slag (174.93g, Zn1.5%, Fe 4.8%, Pb 10.34%, S 82.6%); the hot acid slag is sent to the flotation process to recover lead.

[0060] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 48.98 g of zinc roasted sand was slowly added, and the pH value was controlled to be 2.5 at the end point. Neutralized slag (24.49 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 73.70 g / L, Fe 29.26 g / L, Fe 2+ 0.65g / L), neutralize the slag and return oxygen pressure leaching.

[0061] (4) 120 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 81.18 g / L, Fe 29.85 g / L, Fe 2+ 29.55g / L).

[0062] (5) returning the reduced slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0063] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 81.16 g / L, Fe 1.5 g / L, H2SO4 39 g / L) and iron slag (56.32 g, Zn 0.19%, Fe 56.55%, S 1.0%); wherein, the steam consumption in the iron removal process was 1.08 t steam / t zinc concentrate.

[0064] (6) Iron slag is sent for subsequent resource recovery;

[0065] The iron removal solution is returned to oxygen pressure leaching.

[0066] According to calculation, in the above whole process, the recovery rate of Zn is 98.93%; the recovery rate of iron is 75.53%.

[0067] Comparative Example 1

[0068] Example 1 was repeated, with the main difference being that the amount of waste electrolyte added in step (2) was 0.6 L. The specific steps included:

[0069] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 11.58 g of neutralized slag (Zn 53%, Fe 9%, S1.5%, Pb 5.05%), and 95.04 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.65 m 2 / cm 3 ), 2.2g of lignin, 0.6L of iron removal liquid (65g / L H2SO4, 89.71g / L Zn, 3.2g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 131.5g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 278.64g of oxygen leaching residue (5.5% Zn, 13.3% Fe, 6.12% Pb, 49.86%) and 2L of oxygen leaching liquid (59g / L H2SO4, 165.64g Zn, 6.16g Fe, 4.77g / L Fe) were obtained. 2+ 2.8g / L); the steam consumption of oxygen pressure leaching process is 90kg steam / t zinc concentrate.

[0070] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0071] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0072] (2) 0.6 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 278.64 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (0.6 L, H2SO4 52 g / L, Zn 60.88 g / L, Fe43.24 g / L, Fe 2+ 1.2g / L) and hot acid slag (200.62g, Zn4.2%, Fe 5.7%, Pb 8.50%, S 78.5%); the hot acid slag is sent to the flotation process to recover lead.

[0073] (3) The 0.6 L hot acid solution was reacted at 85° C. for 75 min, during which 23.15 g of zinc roasted sand was slowly added, and the pH value was controlled to be 2.5 at the end point. Neutralized slag (11.58 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (0.6 L, Zn 77.38 g / L, Fe 43.67 g / L, Fe 2+ 1.67g / L), neutralize the slag and return oxygen pressure leaching.

[0074] (4) 108 g of zinc concentrate was added to the 0.6 L neutralization solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (95.04 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (0.6 L, Zn 89.72 g / L, Fe 44.54 g / L, Fe 2+ 44.10g / L).

[0075] (5) returning the reduced slag (95.04 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0076] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (0.6L, Zn 89.71g / L, Fe 3.2g / L, H2SO4 65g / L) and iron slag (45.84g, Zn 0.58%, Fe 56.12%, S 2.4%). The steam consumption in the iron removal process was 1.07t steam / t zinc concentrate.

[0077] (6) Iron slag is sent for subsequent resource recovery;

[0078] The iron removal solution is returned to oxygen pressure leaching.

[0079] According to calculation, in the above whole process, the recovery rate of Zn is 96.51%; the recovery rate of iron is 72.12%.

[0080] The results showed that too little waste electrolyte was added during the hot acid leaching process, resulting in a significant decrease in the zinc leaching rate, which could not reach the target value.

[0081] Example 2

[0082] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0083] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 20.88 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 105.6 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.73 m 2 / cm 3 ), 2.2g of lignin, 0.93L of iron removal liquid (40g / L H2SO4, 84.73g / L Zn, 1.52g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 112.1g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 289.56g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.18% Pb, and 49.86% S) and 2.33L of oxygen leaching liquid (50g / L H2SO4, 156.40g / L Zn, 5.8g / L Fe, and 43g / L S) were obtained. 2+ 2.2g / L); the steam consumption of oxygen pressure leaching process is 78kg steam / t zinc concentrate.

[0084] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0085] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0086] (2) 0.93 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 289.56 g oxygen leaching residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (0.93 L, H2SO4 50 g / L, Zn 56.70 g / L, Fe33.70 g / L, Fe 2+ 0.9g / L) and hot acid slag (182.42g, Zn1.52%, Fe 4.8%, Pb 9.81%, S 82.3%); the hot acid slag is sent to the flotation process to recover lead.

[0087] (3) The 0.93 L hot acid solution was reacted at 85° C. for 75 min, during which 41.76 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (20.88 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (0.93 L, Zn 75.89 g / L, Fe 34.38 g / L, Fe 2+ 0.78g / L), neutralize the slag and return oxygen pressure leaching.

[0088] (4) 120 g of zinc concentrate was added to the 0.93 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (0.93 L, Zn 84.75 g / L, Fe 35.06 g / L, Fe 2+ 34.71g / L).

[0089] (5) returning the reduced slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0090] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (0.93L, Zn 84.73g / L, Fe 1.52g / L, H2SO4 40g / L) and iron slag (55.93g, Zn 0.20%, Fe 56.55%, S 1.2%); wherein, the steam consumption in the iron removal process was 1.06t steam / t zinc concentrate.

[0091] (6) Iron slag is sent for subsequent resource recovery;

[0092] The iron removal solution is returned to oxygen pressure leaching.

[0093] It was calculated that in the above whole process, the recovery rate of Zn was 98.91%; and the recovery rate of iron was 75.34%.

[0094] Example 3

[0095] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0096] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 36.36 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 105.6 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.72 m 2 / cm 3), 2.2g of lignin, 1.4L of iron removal liquid (32g / L H2SO4, 80.44g / L Zn, 1.48g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 107.52g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 298.08g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.27% Pb, and 49.86% S) and 2.8L of oxygen leaching liquid (50g / L H2SO4, 137.10g / L Zn, 5.2g / L Fe, and 43g / L S) were obtained. 2+ 2g / L); the steam consumption of oxygen pressure leaching process is 80kg steam / t zinc concentrate.

[0097] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0098] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0099] (2) 1.4 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 298.08 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.4 L, H2SO4 70 g / L, Zn 52.37 g / L, Fe23.05 g / L, Fe 2+ 0.7g / L) and hot acid slag (172.89g, Zn1.48%, Fe 4.8%, Pb 10.81%, S 82.3%); the hot acid slag is sent to the flotation process to recover lead.

[0100] (3) The 1.4 L hot acid solution was reacted at 85° C. for 75 min, during which 72.73 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (36.36 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.4 L, Zn 74.58 g / L, Fe 23.51 g / L, Fe 2+ 0.65g / L), neutralize the slag and return oxygen pressure leaching.

[0101] (4) 120 g of zinc concentrate was added to the 1.4 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.4 L, Zn 80.46 g / L, Fe 23.98 g / L, Fe 2+ 23.74g / L).

[0102] (5) returning the reduced slag (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0103] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.4 L, Zn 80.44 g / L, Fe 1.48 g / L, H2SO4 32 g / L) and iron slag (57.58 g, Zn 0.19%, Fe 56.55%, S 0.9%). The steam consumption in the iron removal process was 1.10 t steam / t zinc concentrate.

[0104] (6) Iron slag is sent for subsequent resource recovery;

[0105] The iron removal solution is returned to oxygen pressure leaching.

[0106] According to calculation, in the above whole process, the recovery rate of Zn is 98.95%; the recovery rate of iron is 76.08%.

[0107] Comparative Example 2

[0108] Example 1 was repeated, with the main difference being that the amount of waste electrolyte added in step (2) was 1.5 L. The specific steps included:

[0109] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 41.74 g of neutralized slag (Zn 53%, Fe 8.9%, S 1.5%, Pb 5.05%), and 110 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.68 m 2 / cm 3), 2.22g of lignin, 1.5L of iron removal liquid (30g / L H2SO4, 81.40g / L Zn, 1.51g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 92.7g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, 303.46g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.30% Pb, and 49.86% S) and 2.9L of oxygen leaching liquid (45g / L H2SO4, 134.22g / L Zn, 5.04g / L Fe, and 5.70% Pb) were obtained after cooling, pressure reduction, and liquid-solid separation. 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 88kg steam / t zinc concentrate.

[0110] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0111] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0112] (2) 1.5 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 303.46 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.5 L, H2SO4 75 g / L, Zn 51.90 g / L, Fe22.03 g / L, Fe 2+ 0.62g / L) and hot acid slag (179.04g, Zn1.48%, Fe 4.7%, Pb 10.67%, S 82.7%); the hot acid slag is sent to the flotation process to recover lead.

[0113] (3) The 1.5 L hot acid solution was reacted at 85° C. for 75 min, during which 83.49 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (41.74 g, Zn 53%, Fe 8.9%, Pb 5.05%, S 1.5%) and neutralized solution (1.5 L, Zn 75.7 g / L, Fe 22.69 g / L, Fe 2+ 0.58g / L), neutralize the slag and return oxygen pressure leaching.

[0114] (4) 125 g of zinc concentrate was added to the 1.5 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (110 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.5 L, Zn 81.41 g / L, Fe 23.15 g / L, Fe 2+ 22.91g / L).

[0115] (5) returning the reduced slag (110 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0116] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.5 L, Zn 81.40 g / L, Fe 1.51 g / L, H2SO4 30 g / L) and iron slag (59.55 g, Zn 0.19%, Fe 56.51%, S 1.2%). Among them, the steam consumption of the iron removal process was 1.85 t steam / t zinc concentrate.

[0117] (6) Iron slag is sent for subsequent resource recovery;

[0118] The iron removal solution is returned to oxygen pressure leaching.

[0119] It is calculated that in the above whole process, the recovery rate of Zn is 98.92%; the recovery rate of iron is 75.54%.

[0120] The results show that excessive addition of spent electrolyte in the hot acid leaching process will lead to an increase in the amount of roasted sand used for neutralization and excessive steam consumption in the subsequent iron removal process.

[0121] Example 4

[0122] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0123] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 23.27 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 108.24 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.71 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 80.43g / L Zn, 1.52g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 292.33g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.34g / L Zn, 5.6g / L Fe, 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 82kg steam / t zinc concentrate.

[0124] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0125] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0126] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 292.33 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.69 g / L, Fe28.77 g / L, Fe 2+ 0.71g / L) and hot acid slag (175.4g, Zn1.5%, Fe 4.8%, Pb 10.33%, S 82.6%); the hot acid slag is sent to the flotation process to recover lead.

[0127] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 46.53 g of zinc roasted sand was slowly added, and the pH value was controlled at 2 at the end point. Neutralized slag (23.27 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 72.78 g / L, Fe 29.34 g / L, Fe 2+ 0.66g / L), neutralize the slag and return oxygen pressure leaching.

[0128] (4) 123 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (108.2 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 80.45 g / L, Fe 29.93 g / L, Fe 2+ 29.63g / L).

[0129] (5) returning the reduced slag (108.2 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0130] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 80.43 g / L, Fe 1.52 g / L, H2SO4 39 g / L) and iron slag (56.47 g, Zn 0.19%, Fe 56.55%, S 1.1%); wherein, the steam consumption in the iron removal process was 1.08 t steam / t zinc concentrate.

[0131] (6) Iron slag is sent for subsequent resource recovery;

[0132] The iron removal solution is returned to oxygen pressure leaching.

[0133] It is calculated that in the above whole process, the recovery rate of Zn is 98.93%; the recovery rate of iron is 75.52%.

[0134] Example 5

[0135] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0136] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 25.71 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 100.32 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.73 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 81.65g / L Zn, 1.49g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 289.32g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.08g / L Zn, 5.6g / L Fe, 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 79kg steam / t zinc concentrate.

[0137] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0138] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0139] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 289.32 g of oxygen leaching residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.57 g / L, Fe28.47 g / L, Fe 2+ 0.69g / L) and hot acid slag (173.59g, Zn1.5%, Fe 4.8%, Pb 10.34%, S 82.6%); the hot acid slag is sent to the flotation process to recover lead.

[0140] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 51.43 g of zinc roasted sand was slowly added, and the pH value was controlled at 4 at the end point. Neutralized slag (25.71 g, Zn 53%, Fe 9.2%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 74.56 g / L, Fe 28.75 g / L, Fe 2+ 0.63g / L), neutralize the slag and return oxygen pressure leaching.

[0141] (4) 114 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (100.3 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 81.67 g / L, Fe 29.33 g / L, Fe 2+ 29.04g / L).

[0142] (5) returning the reduced slag (100.3 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0143] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 81.65 g / L, Fe 1.49 g / L, H2SO4 39 g / L) and iron slag (55.34 g, Zn 0.19%, Fe 56.55%, S1%). Among them, the steam consumption of the iron removal process was 1.08 t steam / t zinc concentrate.

[0144] (6) Iron slag is sent for subsequent resource recovery;

[0145] The iron removal solution is returned to oxygen pressure leaching.

[0146] It is calculated that in the above whole process, the recovery rate of Zn is 98.92%; the recovery rate of iron is 75.45%.

[0147] Example 6

[0148] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0149] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 25.31 g of neutralized slag (Zn 53%, Fe 9%, S1.5%, Pb 5.05%), and 103.02 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.73 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 81.58g / L Zn, 1.49g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 290.58g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.39g / L Zn, 5.6g / L Fe, 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 78kg steam / t zinc concentrate.

[0150] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0151] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0152] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 290.58 g oxygen leached residue, and the mixture was reacted at 90 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 62 g / L, Zn 54.62 g / L, Fe27.90 g / L, Fe 2+ 0.7g / L) and hot acid slag (180.16g, Zn1.54%, Fe 5.1%, Pb 10.01%, S 82.5%); the hot acid slag is sent to the flotation process to recover lead.

[0153] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 50.61 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (25.31 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 74.29 g / L, Fe 28.45 g / L, Fe 2+ 0.65g / L), neutralize the slag and return oxygen pressure leaching.

[0154] (4) 117.07 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (103.02 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 81.59 g / L, Fe 29.02 g / L, Fe 2+ 28.73g / L).

[0155] (5) returning the reduced slag (103.02 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0156] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 81.58 g / L, Fe 1.49 g / L, H2SO4 39 g / L) and iron slag (54.76 g, Zn 0.19%, Fe 56.54%, S 1.1%); wherein, the steam consumption in the iron removal process was 1.08 t steam / t zinc concentrate.

[0157] (6) Iron slag is sent for subsequent resource recovery;

[0158] The iron removal solution is returned to oxygen pressure leaching.

[0159] It is calculated that in the above whole process, the recovery rate of Zn is 98.92%; the recovery rate of iron is 75.53%.

[0160] Example 7

[0161] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0162] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 23.67 g of neutralized slag (Zn 53%, Fe 9%, S1.5%, Pb 5.05%), and 107.36 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.71 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 81.11g / L Zn, 1.51g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 292.07g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.56g / L Zn, 5.6g / L Fe, 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 81kg steam / t zinc concentrate.

[0163] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0164] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0165] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 292.07 g oxygen leached residue, and the mixture was reacted at 95 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 58 g / L, Zn 55.12 g / L, Fe29.09 g / L, Fe 2+ 0.7g / L) and hot acid slag (172.32g, Zn1.46%, Fe 4.7%, Pb 10.51%, S 82.7%); the hot acid slag is sent to the flotation process to recover lead.

[0166] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 47.35 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (23.67 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 73.52 g / L, Fe 29.67 g / L, Fe 2+ 0.66g / L), neutralize the slag and return oxygen pressure leaching.

[0167] (4) 122 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, liquid-solid separation was performed to obtain reduced slag (107.4 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and reduced solution (1.1 L, Zn 81.13 g / L, Fe 30.27 g / L, Fe 2+ 29.96g / L).

[0168] (5) returning the reduced slag (107.4 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0169] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1L, Zn 81.11g / L, Fe 1.51g / L, H2SO4 39g / L) and iron slag (57.11g, Zn 0.19%, Fe 56.57%, S 1.15%); wherein, the steam consumption in the iron removal process was 1.09t steam / t zinc concentrate.

[0170] (6) Iron slag is sent for subsequent resource recovery;

[0171] The iron removal solution is returned to oxygen pressure leaching.

[0172] According to calculation, in the above whole process, the recovery rate of Zn is 98.97%; the recovery rate of iron is 75.87%.

[0173] Comparative Example 3

[0174] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0175] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 22.04 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 109.12 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.7 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 78.39g / L Zn, 1.5g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 108.08g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 292.14g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.19% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 150.41g / L Zn, 5.6g / L Fe 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 86kg steam / t zinc concentrate.

[0176] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0177] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0178] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 292.14 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.69 g / L, Fe28.75 g / L, Fe 2+ 0.7g / L) and hot acid slag (175.28g, Zn1.5%, Fe 4.8%, Pb 10.32%, S 82.4%); the hot acid slag is sent to the flotation process to recover lead.

[0179] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 44.08 g of zinc roasted sand was slowly added, and the pH value was controlled to be 1 at the end point. Neutralized slag (22.04 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 70.68 g / L, Fe 29.32 g / L, Fe 2+ 0.64g / L), neutralize the slag and return oxygen pressure leaching.

[0180] (4) 124 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (109.12 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 78.41 g / L, Fe 29.91 g / L, Fe 2+ 29.61g / L).

[0181] (5) returning the reduced slag (109.12 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0182] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 78.39 g / L, Fe 1.52 g / L, H2SO4 42 g / L) and iron slag (56.43 g, Zn 0.19%, Fe 56.52%, S 1.23%). Among them, the steam consumption of the iron removal process was 1.09 t steam / t zinc concentrate.

[0183] (6) Iron slag is sent for subsequent resource recovery;

[0184] The iron removal solution is returned to oxygen pressure leaching.

[0185] According to calculation, in the above whole process, the recovery rate of Zn is 98.93%; the recovery rate of iron is 75.46%.

[0186] The results showed that lowering the pH control end point of the neutralization process resulted in the generation of a rotten egg smell during the reduction process, which increased the amount of zinc concentrate used in the reduction process. It had no significant effect on improving the zinc leaching rate and iron resource utilization rate, and the amount of reduced slag required to be returned increased, which was not conducive to reducing the energy consumption of material transportation.

[0187] Comparative Example 4

[0188] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0189] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 26.94 g of neutralized slag (Zn 53%, Fe 9%, S1.5%, Pb 5.05%), and 103.84 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.69 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 81.56g / L Zn, 1.5g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 291.93g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.22% Pb, and 49.86% S) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.84g / L Zn, 5.6g / L Fe, and 43g / L S) were obtained. 2+ 2.1g / L); the steam consumption of oxygen pressure leaching process is 89kg steam / t zinc concentrate.

[0190] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0191] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0192] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 291.93 g oxygen leached residue, and the mixture was reacted at 92 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.68 g / L, Fe28.73 g / L, Fe 2+ 0.69g / L) and hot acid slag (175.16g, Zn1.5%, Fe 4.8%, Pb 10.36%, S 82.5%); the hot acid slag is sent to the flotation process to recover lead.

[0193] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 53.88 g of zinc roasted sand was slowly added, and the pH value was controlled to be 5 at the end point. Neutralized slag (26.94 g, Zn 53%, Fe 12.5%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 74.22 g / L, Fe 25.28 g / L, Fe 2+ 0.61g / L), neutralize the slag and return oxygen pressure leaching.

[0194] (4) 118 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (103.84 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 81.58 g / L, Fe 25.78 g / L, Fe 2+ 25.53g / L).

[0195] (5) returning the reduced slag (103.84 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0196] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 81.56 g / L, Fe 1.51 g / L, H2SO4 39 g / L) and iron slag (48.65 g, Zn 0.19%, Fe 56.53%, S 1.2%). Among them, the steam consumption of the iron removal process was 1.08 t steam / t zinc concentrate.

[0197] (6) Iron slag is sent for subsequent resource recovery;

[0198] The iron removal solution is returned to oxygen pressure leaching.

[0199] According to calculation, in the above whole process, the recovery rate of Zn is 98.93%; the recovery rate of iron is 72.23%.

[0200] The results showed that flocculent iron hydroxide precipitates during the neutralization process, which causes the iron ions in the hot acid solution to disperse, reducing the utilization rate of iron resources. In addition, the iron loss caused by the neutralization process further causes a reduction in the amount of concentrate used for subsequent zinc concentrate reduction, that is, a reduction in the amount of reduction slag produced, that is, a reduction in the amount of reduction slag returned for oxygen pressure leaching, resulting in an increase in steam usage in the oxygen pressure leaching process.

[0201] Comparative Example 5

[0202] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0203] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 25.71 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 95.04 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.67 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (39g / L H2SO4, 77.44g / L Zn, 1.5g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 106.76g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 286.41g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 148.57g / L Zn, 5.4g / L Fe, 2+ 2g / L); the steam consumption of oxygen pressure leaching process is 95kg steam / t zinc concentrate.

[0204] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0205] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0206] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 286.41 g oxygen leached residue, and the mixture was reacted at 85 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 63 g / L, Zn 50.73 g / L, Fe24.75 g / L, Fe 2+ 0.7g / L) and hot acid slag (186.17g, Zn3.82%, Fe5.7%, Pb 9.54%, S 81.2%); the hot acid slag is sent to the flotation process to recover lead.

[0207] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 51.43 g of zinc roasted sand was slowly added, and the pH value was controlled to be 2.5 at the end point. Neutralized slag (25.71 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 70.72 g / L, Fe 25.24 g / L, Fe 2+ 0.63g / L), neutralize the slag and return oxygen pressure leaching.

[0208] (4) 108 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, liquid-solid separation was performed to obtain reduced slag (95.04 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and reduced solution (1.1 L, Zn 77.46 g / L, Fe 25.75 g / L, Fe 2+ 25.49g / L).

[0209] (5) returning the reduced slag (95.04 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0210] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 77.44 g / L, Fe 1.5 g / L, H2SO4 39 g / L) and iron slag (48.58 g, Zn 0.22%, Fe 56.53%, S 1.2%). Among them, the steam consumption of the iron removal process was 1.08 t steam / t zinc concentrate.

[0211] (6) Iron slag is sent for subsequent resource recovery;

[0212] The iron removal solution is returned to oxygen pressure leaching.

[0213] According to calculation, in the above whole process, the recovery rate of Zn is 97.10%; the recovery rate of iron is 73.22%.

[0214] The results show that the reduction of hot acid temperature will lead to a decrease in the recovery rates of Zn and Fe.

[0215] Comparative Example 6

[0216] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0217] (1) 400 g of zinc concentrate (sulfide zinc concentrate), 24.08 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 106.48 g of reduced slag (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.66 m 2 / cm 3), 2.2g of lignin, 1.1L of iron removal liquid (40g / L H2SO4, 81.21g / L Zn, 1.51g / L Fe), and 1.4L of spent electrolyte (160g / L H2SO4, 43g / L Zn) were mixed. At this time, the initial sulfuric acid concentration of the reaction system was 107.2g / L. Oxygen with a purity of 99vol% was introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature was controlled to be 150°C and the total pressure was 1.25MPa. After reacting for 2.5h, after cooling and depressurizing and liquid-solid separation, 291.81g of oxygen leaching residue (5.5% Zn, 13.2% Fe, 6.20% Pb, 49.86%) and 2.5L of oxygen leaching liquid (50g / L H2SO4, 151.54g / L Zn, 5.5g / L Fe, 2+ 2g / L); the steam consumption of oxygen pressure leaching process is 98kg steam / t zinc concentrate.

[0218] The oxygen immersion solution can be sent to subsequent pre-neutralization, purification, electrolysis and casting processes.

[0219] The contents of relevant elements in the zinc concentrate are as follows: Zn 49%, Fe 8%, S29%, Pb 3.24%; the proportion of solid matter with a particle size of -37μm in the zinc concentrate is greater than 97%, and the specific surface area of ​​the zinc concentrate is 3.98m 2 / cm 3 .

[0220] (2) 1.1 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added to the 291.81 g oxygen leached residue, and the mixture was reacted at 97 °C for 2 h, followed by solid-liquid separation to obtain hot acid solution (1.1 L, H2SO4 59 g / L, Zn 54.96 g / L, Fe28.71 g / L, Fe 2+ 0.7g / L) and hot acid slag (175.09g, Zn1.47%, Fe 4.7%, Pb10.34%, S 82.6%); the hot acid slag is sent to the flotation process to recover lead.

[0221] (3) The 1.1 L hot acid solution was reacted at 85° C. for 75 min, during which 48.16 g of zinc roasted sand was slowly added, and the pH value was controlled at 2.5 at the end point. Neutralized slag (24.08 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (1.1 L, Zn 73.68 g / L, Fe 29.29 g / L, Fe 2+ 0.64g / L), neutralize the slag and return oxygen pressure leaching.

[0222] (4) 121 g of zinc concentrate was added to the 1.1 L neutralized solution, and the mixture was reacted at 75° C. for 300 min. Then, a reduction slag (106.48 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction solution (1.1 L, Zn 81.23 g / L, Fe 29.87 g / L, Fe 2+ 29.57g / L).

[0223] (5) returning the reduced slag (106.48 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) to oxygen pressure leaching;

[0224] The reducing liquid was added into an autoclave, and oxygen with a purity of 99 vol% was introduced at 190°C. The total pressure was controlled to be 1.8 MPa. After reacting for 300 minutes, the temperature and pressure were reduced to obtain iron removal liquid (1.1 L, Zn 81.21 g / L, Fe 1.51 g / L, H2SO4 40 g / L) and iron slag (56.37 g, Zn 0.19%, Fe 56.54%, S 1.21%). Among them, the steam consumption of the iron removal process was 1.08 t steam / t zinc concentrate.

[0225] (6) Iron slag is sent for subsequent resource recovery;

[0226] The iron removal solution is returned to oxygen pressure leaching.

[0227] It is calculated that in the above whole process, the recovery rate of Zn is 98.95%; the recovery rate of iron is 75.62%.

[0228] The results show that increasing the hot acid temperature will increase energy consumption, and will not significantly improve the recovery rates of Zn and Fe.

[0229] The contents explained in the above embodiments should be understood as these embodiments are only used to more clearly illustrate the present invention, and are not used to limit the scope of the present invention. After reading the present invention, various equivalent forms of modifications to the present invention by those skilled in the art all fall within the scope defined by the claims attached to this application.

Claims

1. A method for extracting iron by zinc-oxygen pressure leaching, characterized in that: The steps include: S1. After oxygen pressure leaching of the zinc concentrate to be treated, solid-liquid separation is performed to obtain oxygen leaching slag and oxygen leaching liquid; Wherein, the zinc concentrate is zinc sulfide concentrate; S2, after hot acid leaching the oxygen leaching residue at 90-95°C, solid-liquid separation is performed to obtain hot acid slag and Fe-rich 3+ of hot acid; S3, after neutralizing the hot acid liquid, separate the solid and the liquid to obtain a neutralized slag and a neutralized liquid with a pH value of 2 to 4; S4, returning the neutralized slag to the oxygen pressure leaching process in S1; After the neutralized liquid is reduced by a reducing agent, solid-liquid separation is performed to obtain a reduced liquid and reduced slag; Wherein, the reducing agent is ZnS and / or zinc concentrate; S5, after removing iron from the reducing liquid by the hematite method, separating the solid from the liquid to obtain iron removal liquid and iron slag; Returning the reduced slag to the oxygen pressure leaching step S1; S6, returning the iron removal solution to the oxygen pressure leaching process in S1.

2. The method according to claim 1, characterized in that In S1, when oxygen pressure leaching is performed, oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced, and a first solution containing sulfuric acid is added, and the initial acid concentration of the reaction system is controlled to be 90-125 g / L, the reaction temperature is 135-158° C., the total pressure is 1.15-1.35 MPa, and the reaction time is 2-3.5 h; preferably, when oxygen pressure leaching is performed, a surfactant is added to the reaction system, wherein the amount of the surfactant added is 3-6‰ of the zinc concentrate; preferably, the surfactant is lignin; Preferably, the concentration of sulfuric acid in the first solution is 140-180 g / L, and more preferably, the first solution is waste electrolyte.

3. The method according to claim 2, characterized in that In S2, when hot acid leaching is performed, the leaching temperature is controlled to be 91 to 94°C, preferably 92 to 93°C.

4. The method according to claim 2, characterized in that: The oxygen leaching residue is mixed with a second solution containing sulfuric acid to perform 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 oxygen leaching residue to the second solution is 1 g: 2-6 mL, preferably 1 g: 3-5 mL; more preferably, the second solution is a waste electrolyte; Preferably, the amount of the second solution added in S2 is 38-51 vol%, preferably 40-50 vol%, and more preferably 42-48 vol% of the total amount of the first solution in S1 and the second solution in S2.

5. The method according to claim 1, characterized in that In S2, when hot acid leaching is performed, the initial acid concentration of the reaction system is 140 to 180 g / L.

6. The method according to claim 1, characterized in that In S3, the hot acid solution is neutralized by using zinc oxide and / or zinc roasted sand; Preferably, during the neutralization treatment, the reaction temperature is controlled to be 80-90°C; Preferably, during the neutralization treatment, the reaction time is controlled to be 60-120 min.

7. The method according to claim 1, characterized in that In S4, when the reduction treatment is performed, the amount of the reducing agent added is such that Fe 3+ Completely converted to Fe 2+ 1.5-2.5 times, preferably 1.75-2.25 times, the amount of the required reducing agent; Preferably, during the reduction treatment, the reaction temperature is controlled to be 70-80° C. and the reaction time is 4-6 hours.

8. The method according to claim 1, characterized in that In S5, when the hematite method is used for iron removal, an oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced into the reducing liquid, and the reaction temperature is controlled to be 180-200° C. and the total pressure is 1.6-2.0 MPa; preferably, the reaction time is 4-6 h.

9. The method according to any one of claims 1 to 8, characterized in that: In the zinc concentrate, the content of Zn is 40-60%, the content of Fe is 4-12%, and the content of S is 20-40%; Preferably, in the zinc concentrate, the Zn content is 48-51%, the Fe content is 6-9%, and the S content is 28-32%.

10. The method according to any one of claims 1 to 8, characterized in that: In the zinc concentrate, the content of solid matter with a particle size of -37 μm is greater than 95 wt %.

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