A method for zinc extraction by zinc oxygen pressure leaching

Abstract

The application relates to a method for zinc oxygen pressure leaching and iron extraction, which comprises the following steps: after oxygen pressure leaching of a zinc concentrate to be treated, solid-liquid separation is carried out to obtain oxygen leaching residue and oxygen leaching liquid; after hot acid leaching of the oxygen leaching residue, solid-liquid separation is carried out to obtain hot acid residue and hot acid liquid rich in Fe 3+ ; after neutralization treatment of the hot acid liquid, solid-liquid separation is carried out to obtain neutralization residue and neutralization liquid with a pH value of 2-4; the neutralization residue is returned to the oxygen pressure leaching procedure in S1; after reduction treatment of the neutralization liquid by a reducing agent, solid-liquid separation is carried out to obtain reduction liquid and reduction residue; wherein the reducing agent is ZnS and / or zinc concentrate; after hematite method iron removal of the reduction liquid, solid-liquid separation is carried out to obtain iron removal liquid and iron residue; the reduction residue is returned to the oxygen pressure leaching procedure in S1; and the iron removal liquid is returned to the oxygen pressure leaching procedure in S1. The application can realize iron resourceization of the zinc oxygen pressure leaching system while reducing the loss of valuable elements such as zinc, and the Zn recovery rate can reach more than 98.9%.

Description

TECHNICAL FIELD

[0001] The present application relates to a method for zinc extraction from zinc oxide pressure leaching, and belongs to the field of hydrometallurgy. BACKGROUND

[0002] With the continuous decrease of high-zinc and low-iron zinc-containing minerals, zinc smelting enterprises are increasingly choosing zinc materials with lower zinc grade 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 treatment load, high valuable metal content taken away by 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, and in the iron removal process, either in-situ iron precipitation or ex-situ goethite iron removal is used. The main problem is that the volume of zinc-containing solution that needs to be removed is large, which leads to increased steam consumption, increased neutralizing agent consumption, and increased iron slag volume. Moreover, the low iron content in the obtained iron slag leads to poor comprehensive recovery of iron, and the comprehensive benefits of zinc smelting enterprises need to be improved. The current zinc smelting enterprises 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 staged leaching process, the one-stage oxygen pressure direct leaching process mainly has high sulfuric acid concentration and iron concentration in the oxygen pressure leaching solution, which brings great pressure to the subsequent acid reduction and iron removal. The existing technology usually considers recycling or recycling the iron in the leaching solution. Due to the large liquid-solid ratio and large volume of iron-containing solution, it will cause problems such as high energy consumption.

[0003] Therefore, it is urgent to develop more advanced zinc oxygen pressure leaching process technology to achieve efficient zinc leaching while reducing the volume of liquid required for iron treatment and improving the economic benefits of the zinc smelting system.

[0004] The Chinese patent application specification CN118345248A discloses a method for opening circuit iron from a high-iron sphalerite oxygen pressure leaching system, first finely grinding the high-iron sphalerite, then performing one-stage oxygen pressure leaching to obtain one-stage leaching slurry, then thickening the one-stage leaching slurry to obtain thickening underflow and supernatant; then performing rough separation on the thickening underflow to obtain once-roughed concentrate and tailings, performing one-stage scavenging on the once-roughed tailings to obtain one-stage scavenged concentrate and tailings, and performing two-stage scavenging on the one-stage scavenged tailings to obtain two-stage scavenged concentrate and final iron-containing tailings; mixing the roughed concentrate, the one-stage scavenged concentrate and the two-stage scavenged concentrate, and then performing one-stage mixed cleaning to obtain final zinc-sulfur concentrate and one-stage cleaning tailings, and then returning the one-stage cleaning tailings to the roughing operation to form a closed circuit, and the thickening supernatant of the one-stage oxygen pressure leaching slurry enters a zinc smelting and iron removal process. Although the technology solves the problem that the existing iron removal method is not applicable to the opening circuit iron in the oxygen pressure leaching system, it needs to rely on a relatively complex flotation process, and the flotation concentration and the flotation slurry temperature need to be strictly controlled to strengthen the separation of sulfur, zinc and iron, which further increases the operation difficulty; in addition, the final iron-containing tailings obtained have iron content of more than 30%, zinc content of more than 2.5% and sulfur content of more than 17%, the iron grade is low, and the loss rates of zinc and sulfur are high, so the comprehensive utilization value is low. SUMMARY

[0005] In view of the deficiencies of the prior art, the purpose of the present application is to provide a zinc oxygen pressure leaching zinc iron extraction method, based on a one-stage oxygen pressure leaching method with shorter process, to reduce the loss of valuable elements such as zinc while recycling the iron in the zinc oxygen pressure leaching system.

[0006] To solve the above technical problems, the technical solution of the present application is as follows:

[0007] A zinc oxygen pressure leaching zinc iron extraction method, 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 residue and oxygen leaching liquid;

[0009] The zinc concentrate is a zinc sulfide concentrate;

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

[0011] S3, after neutralization treatment of the hot acid liquid, solid-liquid separation is performed to obtain neutralization residue and neutralization liquid with pH value of 2-4;

[0012] S4, the neutralization residue is returned to the oxygen pressure leaching process in S1;

[0013] After reduction treatment of the neutralization liquid with a reducing agent, solid-liquid separation is performed to obtain reduction liquid and reduction residue;

[0014] The reducing agent is ZnS and / or zinc concentrate;

[0015] S5, after removing iron from the reducing solution by the hematite process, solid-liquid separation is performed to obtain an iron-removed solution and an iron residue;

[0016] The reducing residue is returned to the oxygen pressure leaching process of S1.

[0017] S6, the iron-removed solution is returned to the oxygen pressure leaching process of S1.

[0018] In this way, the zinc concentrate is mixed with the returned neutralizing residue, reducing residue, iron-removed solution and other substances, and after oxygen pressure leaching treatment, oxygen leaching residue and oxygen leaching solution are obtained; the oxygen leaching solution can be sent to the downstream side for pre-neutralization, purification, electrolysis, smelting and other processes, and the oxygen leaching residue is further treated by hot acid leaching to fully leach the valuable elements such as iron and zinc in the oxygen leaching residue, and then the hot acid solution is neutralized to prepare for the next reduction treatment, preventing the generation of toxic gases such as H2S during the reduction treatment using ZnS and / or zinc concentrate as the reducing agent, and allowing the sulfur element to exist in the form of elemental S in the reducing residue. After reduction, the reducing solution is removed by the hematite process to obtain high-grade iron residue and iron-removed solution, and the iron residue can be sold or further processed to realize the resource utilization of iron in the zinc oxygen pressure leaching system; the iron-removed solution is returned to the oxygen pressure leaching process of 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 solution and enter the subsequent process, effectively reducing the loss of valuable metals such as zinc. At the same time, the reducing residue is returned to the oxygen pressure leaching process, and the reducing residue contains elemental S and other active substances such as ZnS that do not participate in the reaction after reduction, and due to the increase in specific surface area and activity, the reactions (1) and (6) are more easily carried out, and heat is stably released to maintain the oxygen pressure leaching reaction, which helps to reduce the consumption of external heat such as steam, and the steam consumption can be reduced to less than 85 kg of steam per ton of zinc concentrate.

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

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

[0021] Preferably, when oxygen pressure leaching is performed, a surfactant is added to the reaction system, wherein the addition amount of the surfactant 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, more preferably 150-170 g / L; more preferably, the first solution is waste electrolyte, thereby effectively disposing of the waste electrolyte in the zinc smelting process and avoiding the introduction of other impurities.

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

[0024] Further, the oxygen leaching residue is mixed with the second solution containing sulfuric acid to perform hot acid leaching; more preferably, the concentration of sulfuric acid in the second solution is 140-180 g / L, preferably 145-175 g / L, more preferably 150-170 g / L; 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 waste electrolyte, thereby effectively disposing of the waste electrolyte in the zinc smelting process and avoiding the introduction of other impurities.

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

[0026] Further, 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, preferably 155-165 g / L.

[0027] Further, in S3, the hot acid liquor is subjected to neutralization treatment by using zinc oxide and / or zinc calcine, so that the neutralization treatment can be realized and the introduction of other elements can be avoided.

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

[0029] Preferably, when the neutralization treatment is performed, the reaction temperature is controlled to be 80-90℃.

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

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

[0032] Preferably, when the reduction treatment is performed, the reaction temperature is controlled to be 70-80℃ and the reaction time is controlled to be 4-6h.

[0033] Further, in S5, when the hematite method is used to remove iron, oxygen gas with an oxygen concentration of ≥50vol% is introduced into the reduction liquor, the reaction temperature is controlled to be 180-200℃, preferably 185-195℃, the total pressure is controlled to be 1.6-2.0MPa, preferably 1.7-1.9MPa; and preferably, the reaction time is 4-6h.

[0034] Further, 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, in the zinc concentrate, the content of Pb is 1.5-5%.

[0036] The main chemical reaction equations that the present application can involve 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] Further, the content of solid phase with particle size of -37 μm in the zinc concentrate is greater than 95 wt%.

[0047] Further, 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 application has the following beneficial effects:

[0049] (1) The method of the present application can realize the opening of iron resources in the zinc oxygen pressure leaching system while reducing the loss of valuable elements such as zinc, and the Zn recovery rate can reach more than 98.9%, and the process is shorter, only one stage of oxygen pressure leaching is needed to meet the efficient recovery of valuable metals such as Zn.

[0050] (2) The iron and zinc in the oxygen leaching slag are leached by hot acid, which ensures the efficient leaching of zinc in the whole process, and the volume of hot acid produced by hot acid leaching is small, which reduces the volume of subsequent iron removal solution and the heat consumption required for the subsequent hematite process for iron removal, thereby reducing the cost of iron removal. The iron-containing solution is subjected to high-temperature high-pressure efficient iron removal, which realizes the technical effect of iron resource utilization, the iron grade of the obtained iron slag is higher than 56%, the impurity content is low, the iron resource utilization rate is high, which is more than 75%, and the content of zinc in the iron slag is ≤0.19%, the content of S is ≤1.2%, the loss of valuable elements such as zinc and sulfur in the zinc smelting system is small.

[0051] (3) The zinc sulfide and / or zinc concentrate are used to reduce the neutralization liquid, and the obtained reduction slag is returned to the oxygen pressure leaching process, which makes the oxygen pressure leaching reaction easier to proceed, which helps to reduce the consumption of external heat, that is, to save the consumption of steam for providing external heat in the oxygen pressure leaching process. BRIEF DESCRIPTION OF DRAWINGS

[0052] Figure 1 is a flow chart of a zinc oxygen pressure leaching zinc extraction iron process of the present application. DETAILED DESCRIPTION

[0053] The present application will be described in detail below with examples. It should be noted that the examples in the present application and the features in the examples can be combined with each other without conflict. If not specifically stated, the relevant percentages refer to mass percentages.

[0054] Example 1

[0055] Referring to Figure 1 , the zinc oxygen pressure leaching zinc extraction iron method of the present embodiment comprises the following steps:

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

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

[0058] Among them, the content of the relevant elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0059] (2) In the 291.55 g oxygen leaching residue, 1.1 L of waste electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92°C for 2 h, solid-liquid separation was performed to obtain hot acid liquor (1.1 L, H2SO4 60 g / L, Zn 54.66 g / L, Fe 28.69 g / L, Fe 2+ 0.7 g / L) and hot acid residue (174.93 g, Zn 1.5%, Fe 4.8%, Pb 10.34%, S 82.6%); the hot acid residue was sent to a flotation process to recover lead.

[0060] (3) To the 1.1 L hot acid liquor, 48.98 g of zinc calcine was slowly added at 85°C for 75 min, and the pH was controlled at 2.5 at the end point, and then liquid-solid separation was performed to obtain neutralization residue (24.49 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization liquor (1.1 L, Zn 73.70 g / L, Fe 29.26 g / L, Fe 2+ 0.65 g / L), and the neutralization residue was returned to oxygen pressure leaching.

[0061] (4) To the 1.1 L neutralization liquor, 120 g of zinc concentrate was added, and after reaction at 75°C for 300 min, liquid-solid separation was performed to obtain reduction residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and reduction liquor (1.1 L, Zn 81.18 g / L, Fe 29.85 g / L, Fe 2+ 29.55 g / L).

[0062] (5) The reduction residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to oxygen pressure leaching;

[0063] The reduction liquor was added to an autoclave, and after reaction at 190°C for 300 min by introducing oxygen with a purity of 99 vol% and controlling the total pressure at 1.8 MPa, the temperature and pressure were reduced to obtain iron removal liquor (1.1 L, Zn 81.16 g / L, Fe 1.5 g / L, H2SO4 39 g / L) and iron residue (56.32 g, Zn 0.19%, Fe 56.55%, S 1.0%); wherein the steam consumption of the iron removal process was 1.08 t of steam / t of zinc concentrate.

[0064] (6) The iron residue was sent to subsequent resource recovery;

[0065] The iron removal liquor was returned to oxygen pressure leaching.

[0066] It was calculated that in the above entire process, the recovery rate of Zn was 98.93%, and the recovery rate of Fe was 75.53%.

[0067] Comparative Example 1

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

[0069] (1) 400 g of zinc concentrate (zinc sulfide concentrate), 11.58 g of neutralization residue (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), 95.04 g of reduction residue (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.65 m 2 / cm 3 ), 2.2 g of lignin, 0.6 L of iron removal solution (H2SO4 65 g / L, Zn 89.71 g / L, Fe 3.2 g / L), and 1.4 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) are mixed. At this time, the initial sulfuric acid concentration of the reaction system is 131.5 g / L. Pure oxygen with a purity of 99 vol% is introduced for oxygen pressure leaching. The oxygen pressure leaching reaction temperature is controlled at 150°C, and the total pressure is 1.25 MPa. After 2.5 h of reaction, after cooling and decompression, and liquid-solid separation, 278.64 g of oxygen leaching residue (Zn 5.5%, Fe 13.3%, Pb 6.12%, S 49.86%) and 2 L of oxygen leaching solution (H2SO4 59 g / L, Zn 165.64 g / L, Fe 6.16 g / L, Fe 2+ 2.8 g / L) are obtained. The steam consumption of the oxygen pressure leaching process is 90 kg of steam per ton of zinc concentrate.

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

[0071] Among them, the content of related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of -37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0072] (2) In the 278.64 g of oxygen leaching residue, 0.6 L of spent electrolyte (H2SO4 160 g / L, Zn 43 g / L) is added, and after 2 h of reaction at 92°C, solid-liquid separation is performed to obtain hot acid solution (0.6 L, H2SO4 52 g / L, Zn 60.88 g / L, Fe 43.24 g / L, Fe 2+ 1.2 g / L) and hot acid residue (200.62 g, Zn 4.2%, Fe 5.7%, Pb 8.50%, S 78.5%); the hot acid residue is sent to the flotation process to recover lead.

[0073] (3) to the 0.6L hot acid solution at 85°C, 23.15g zinc calcine was slowly added for 75min, the end point pH was controlled at 2.5, and then liquid-solid separation was performed to obtain neutralized residue (11.58g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized solution (0.6L, Zn 77.38g / L, Fe 43.67g / L, Fe 2+ 1.67g / L), and the neutralized residue was subjected to pressure leaching.

[0074] (4) to the 0.6L neutralized solution, 108g zinc concentrate was added, and then the reaction was performed at 75°C for 300min, and then liquid-solid separation was performed to obtain reduced residue (95.04g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and reduced solution (0.6L, Zn 89.72g / L, Fe 44.54g / L, Fe 2+ 44.10g / L).

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

[0076] the reduced solution was added into an autoclave, 99vol% pure oxygen was introduced, the total pressure was controlled at 1.8MPa, and then the reaction was performed at 190°C for 300min, and then the temperature and pressure were reduced to obtain iron-removed solution (0.6L, Zn 89.71g / L, Fe 3.2g / L, H2SO4 65g / L) and iron residue (45.84g, Zn 0.58%, Fe 56.12%, S 2.4%); wherein, the steam consumption of the iron-removing process was 1.07t steam / t zinc concentrate.

[0077] (6) the iron residue was sent to subsequent resource recovery;

[0078] the iron-removed solution was returned to pressure leaching.

[0079] It was calculated that the recovery rate of Zn was 96.51% and the recovery rate of Fe was 72.12% in the whole process.

[0080] The results showed that the addition of waste electrolyte in the hot acid leaching process was too small, which led to a significant decrease in zinc leaching rate and could not reach the target value.

[0081] Example 2

[0082] The zinc pressure leaching and iron removal method of the zinc in the example comprises the following steps:

[0083] (1) 400 g of zinc concentrate (zinc sulfide concentrate), 20.88 g of neutralized slag (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), 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 ), 2.2 g of lignin, 0.93 L of iron removal solution (H2SO4 40 g / L, Zn 84.73 g / L, Fe 1.52 g / L), and 1.4 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) are mixed, at this time the initial sulfuric acid concentration of the reaction system is 112.1 g / L, pure oxygen with a purity of 99 vol% is introduced for oxygen pressure leaching, the oxygen pressure leaching reaction temperature is controlled at 150°C, the total pressure is 1.25 MPa, after 2.5 h of reaction, after cooling and pressure reduction and liquid-solid separation, 289.56 g of oxygen leaching slag (Zn 5.5%, Fe 13.2%, Pb 6.18%, S 49.86%) and 2.33 L of oxygen leaching solution (H2SO4 50 g / L, Zn 156.40 g / L, Fe 5.8 g / L, Fe 2+ 2.2 g / L) are obtained; the steam consumption of the oxygen pressure leaching process is 78 kg of steam / t of zinc concentrate.

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

[0085] Among them, the content of the relevant elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0086] (2) In the 289.56 g of oxygen leaching slag, 0.93 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) is added, and after 2 h of reaction at 92°C, solid-liquid separation is performed to obtain hot acid solution (0.93 L, H2SO4 50 g / L, Zn 56.70 g / L, Fe 33.70 g / L, Fe 2+ 0.9 g / L) and hot acid slag (182.42 g, Zn 1.52%, Fe 4.8%, Pb 9.81%, S 82.3%); the hot acid slag is sent to the flotation process to recover lead.

[0087] (3) to the 0.93 L hot acid liquor at 85°C, slowly add zinc calcine 41.76 g, control the pH at 2.5 at the end point, separate the liquid and solid to obtain neutralized residue (20.88 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralized liquor (0.93 L, Zn 75.89 g / L, Fe 34.38 g / L, Fe 2+ 0.78 g / L), and the neutralized residue is returned to the pressure oxidation leaching.

[0088] (4) to the 0.93 L neutralized liquor, add zinc concentrate 120 g, react at 75°C for 300 min, then separate the liquid and solid to obtain reduced residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and reduced liquor (0.93 L, Zn 84.75 g / L, Fe 35.06 g / L, Fe 2+ 34.71 g / L).

[0089] (5) the reduced residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) is returned to the pressure oxidation leaching;

[0090] the reduced liquor is added to an autoclave, oxygen with a purity of 99 vol% is introduced, the total pressure is controlled at 1.8 MPa, and after reacting for 300 min, the temperature and pressure are reduced to obtain iron-removed liquor (0.93 L, Zn 84.73 g / L, Fe 1.52 g / L, H2SO4 40 g / L) and iron residue (55.93 g, Zn 0.20%, Fe 56.55%, S 1.2%); wherein the steam consumption of the iron removal process is 1.06 t of steam per t of zinc concentrate.

[0091] (6) the iron residue is sent to subsequent resource recovery;

[0092] the iron-removed liquor is returned to the pressure oxidation leaching.

[0093] It is calculated that in the above entire process, the recovery rate of Zn is 98.91%, and the recovery rate of Fe is 75.34%.

[0094] Example 3

[0095] The zinc pressure oxidation leaching and iron removal method of the present example comprises the following steps:

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

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

[0098] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0099] (2) In the 298.08 g of oxygen leaching residue, 1.4 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.4 L, H2SO4 70 g / L, Zn 52.37 g / L, Fe 23.05 g / L, Fe 2+ 0.7 g / L) and hot acid residue (172.89 g, Zn 1.48%, Fe 4.8%, Pb 10.81%, S 82.3%); the hot acid residue was sent to a flotation process to recover lead.

[0100] (3) To the 1.4 L of hot acid solution, 72.73 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (36.36 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.4 L, Zn 74.58 g / L, Fe 23.51 g / L, Fe 2+ 0.65 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0101] (4) To the 1.4 L neutralized liquid, 120 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.4 L, Zn 80.46 g / L, Fe 23.98 g / L, Fe 2+ 23.74 g / L).

[0102] (5) The reduction residue (105.6 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0103] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa. After cooling and pressure reduction, a deironing liquid (1.4 L, Zn 80.44 g / L, Fe 1.48 g / L, H2SO4 32 g / L) and an iron residue (57.58 g, Zn 0.19%, Fe 56.55%, S 0.9%) were obtained; wherein the steam consumption of the deironing process was 1.10 t of steam / t of zinc concentrate.

[0104] (6) The iron residue was sent to subsequent resource recovery;

[0105] The deironing liquid was returned to the oxygen pressure leaching.

[0106] It was calculated that in the above entire process, the recovery rate of Zn was 98.95%, and the recovery rate of Fe was 76.08%.

[0107] Comparative Example 2

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

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

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

[0111] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0112] (2) In the 303.46 g of oxygen leaching residue, 1.5 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) is added, and after reaction at 92 ℃ for 2 h, solid-liquid separation is performed to obtain hot acid solution (1.5 L, H2SO4 75 g / L, Zn 51.90 g / L, Fe 22.03 g / L, Fe 2+ 0.62 g / L) and hot acid residue (179.04 g, Zn 1.48%, Fe 4.7%, Pb 10.67%, S 82.7%); the hot acid residue is sent to a flotation process to recover lead.

[0113] (3) To the 1.5 L of hot acid solution, 83.49 g of zinc calcine is slowly added at 85 ℃ for 75 min, and the pH is controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (41.74 g, Zn 53%, Fe 8.9%, Pb 5.05%, S 1.5%) and neutralization solution (1.5 L, Zn 75.7 g / L, Fe 22.69 g / L, Fe 2+ 0.58 g / L) are obtained, and the neutralization residue is returned to the oxygen pressure leaching process.

[0114] (4) to the 1.5 L neutralization liquid, 125 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (110 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.5 L, Zn 81.41 g / L, Fe 23.15 g / L, Fe 2+ 22.91 g / L).

[0115] (5) the reduction residue (110 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0116] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min with the introduction of oxygen with a purity of 99 vol%, the total pressure was controlled at 1.8 MPa, and after cooling and pressure reduction, a deironing liquid (1.5 L, Zn 81.40 g / L, Fe 1.51 g / L, H2SO4 30 g / L) and an iron residue (59.55 g, Zn 0.19%, Fe 56.51%, S 1.2%) were obtained; wherein the steam consumption of the deironing process was 1.85 t of steam / t of zinc concentrate.

[0117] (6) the iron residue was sent to subsequent resource recovery;

[0118] The deironing liquid was returned to the oxygen pressure leaching.

[0119] It was calculated that in the above entire process, the recovery rate of Zn was 98.92%, and the recovery rate of Fe was 75.54%.

[0120] The results show that too much waste electrolyte is added in the hot acid leaching process, which leads to an increase in the amount of calcine used for neutralization and an excessive steam consumption in the subsequent deironing process.

[0121] Example 4

[0122] The zinc oxygen pressure leaching and deironing method of the present embodiment includes the following steps:

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

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

[0125] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0126] (2) In the 292.33 g of oxygen leaching residue, 1.1 L of waste electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.69 g / L, Fe 28.77 g / L, Fe 2+ 0.71 g / L) and hot acid residue (175.4 g, Zn 1.5%, Fe 4.8%, Pb 10.33%, S 82.6%); the hot acid residue was sent to a flotation process to recover lead.

[0127] (3) To the 1.1 L of hot acid solution, 46.53 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2 at the end point, and after liquid-solid separation, neutralization residue (23.27 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 72.78 g / L, Fe 29.34 g / L, Fe 2+ 0.66 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0128] (4) to the 1.1 L of the neutralized liquid, 123 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (108.2 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.1 L, Zn 80.45 g / L, Fe 29.93 g / L, Fe 2+ 29.63 g / L).

[0129] (5) the reduction residue (108.2 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0130] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa, and after cooling and pressure reduction, a deironing liquid (1.1 L, Zn 80.43 g / L, Fe 1.52 g / L, H2SO4 39 g / L) and an iron residue (56.47 g, Zn 0.19%, Fe 56.55%, S 1.1%) were obtained; wherein the steam consumption of the deironing process was 1.08 t of steam / t of zinc concentrate.

[0131] (6) the iron residue was sent to subsequent resource recovery;

[0132] The deironing liquid was returned to the oxygen pressure leaching.

[0133] It was calculated that in the above entire process, the recovery rate of Zn was 98.93%, and the recovery rate of Fe was 75.52%.

[0134] Example 5

[0135] The zinc oxygen pressure leaching and deironing method of the present embodiment comprises the following steps:

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

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

[0138] Among them, the content of related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0139] (2) In the 289.32 g of oxygen leaching residue, 1.1 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.57 g / L, Fe 28.47 g / L, Fe 2+ 0.69 g / L) and hot acid residue (173.59 g, Zn 1.5%, Fe 4.8%, Pb 10.34%, S 82.6%); the hot acid residue was sent to a flotation process to recover lead.

[0140] (3) To the 1.1 L of hot acid solution, 51.43 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 4 at the end point, and after liquid-solid separation, neutralization residue (25.71 g, Zn 53%, Fe 9.2%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 74.56 g / L, Fe 28.75 g / L, Fe 2+ 0.63 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0141] (4) To the 1.1 L neutralized liquid, 114 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (100.3 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.1 L, Zn 81.67 g / L, Fe 29.33 g / L, Fe 2+ 29.04 g / L).

[0142] (5) The reduction residue (100.3 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0143] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa. After cooling and pressure reduction, a deironing liquid (1.1 L, Zn 81.65 g / L, Fe 1.49 g / L, H2SO4 39 g / L) and an iron residue (55.34 g, Zn 0.19%, Fe 56.55%, S 1%) were obtained; wherein the steam consumption of the deironing process was 1.08 t of steam / t of zinc concentrate.

[0144] (6) The iron residue was sent to subsequent resource recovery;

[0145] The deironing liquid was returned to the oxygen pressure leaching.

[0146] It was calculated that in the above entire process, the recovery rate of Zn was 98.92%, and the recovery rate of Fe was 75.45%.

[0147] Example 6

[0148] The zinc oxygen pressure leaching and deironing method of the present embodiment comprises the following steps:

[0149] (1) 400 g of zinc concentrate (zinc sulfide concentrate), 25.31 g of neutralized residue (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 103.02 g of reduction residue (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.73 m 2 / cm 3), lignin 2.2 g, iron-removing solution 1.1 L (H2SO4 39 g / L, Zn 81.58 g / L, Fe 1.49 g / L), waste electrolyte solution 1.4 L (H2SO4 160 g / L, Zn 43 g / L) were mixed, at this time, the initial sulfuric acid concentration of the reaction system was 106.76 g / L, pure oxygen with a purity of 99 vol% was introduced for oxygen pressure leaching, the oxygen pressure leaching reaction temperature was controlled at 150 ℃, the total pressure was 1.25 MPa, after reaction for 2.5 h, after cooling and pressure reduction and liquid-solid separation, oxygen leaching residue 290.58 g (Zn 5.5%, Fe 13.2%, Pb 6.20%, S 49.86%) and oxygen leaching solution 2.5 L (H2SO4 50 g / L, Zn 151.39 g / L, Fe 5.6 g / L, Fe 2+ 2.1 g / L) were obtained; the steam consumption of the oxygen pressure leaching process was 78 kg of steam / t of zinc concentrate.

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

[0151] Among them, the content of related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0152] (2) In the 290.58 g of oxygen leaching residue, 1.1 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 90 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 62 g / L, Zn 54.62 g / L, Fe 27.90 g / L, Fe 2+ 0.7 g / L) and hot acid residue (180.16 g, Zn 1.54%, Fe 5.1%, Pb 10.01%, S 82.5%); the hot acid residue was sent to a flotation process to recover lead.

[0153] (3) To the 1.1 L of hot acid solution, 50.61 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (25.31 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 74.29 g / L, Fe 28.45 g / L, Fe 2+ 0.65 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0154] (4) to the 1.1 L neutralized liquid, 117.07 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (103.02 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.1 L, Zn 81.59 g / L, Fe 29.02 g / L, Fe 2+ 28.73 g / L).

[0155] (5) the reduction residue (103.02 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0156] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa, and after cooling and pressure reduction, a deironing liquid (1.1 L, Zn 81.58 g / L, Fe 1.49 g / L, H2SO4 39 g / L) and an iron residue (54.76 g, Zn 0.19%, Fe 56.54%, S 1.1%) were obtained; wherein the steam consumption of the deironing process was 1.08 t of steam / t of zinc concentrate.

[0157] (6) the iron residue was sent to subsequent resource recovery;

[0158] The deironing liquid was returned to the oxygen pressure leaching.

[0159] It was calculated that in the above entire process, the recovery rate of Zn was 98.92%, and the recovery rate of Fe was 75.53%.

[0160] Example 7

[0161] The zinc oxygen pressure leaching and deironing method of the present embodiment comprises the following steps:

[0162] (1) 400 g of zinc concentrate (zinc sulfide concentrate), 23.67 g of neutralized residue (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 107.36 g of reduction residue (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.71 m 2 / cm 3), lignin 2.2 g, iron-removing solution 1.1 L (H2SO4 39 g / L, Zn 81.11 g / L, Fe 1.51 g / L), waste electrolyte 1.4 L (H2SO4 160 g / L, Zn 43 g / L) were mixed, at this time, the initial sulfuric acid concentration of the reaction system was 106.76 g / L, pure oxygen with a purity of 99 vol% was introduced for oxygen pressure leaching, the oxygen pressure leaching reaction temperature was controlled at 150 ℃, the total pressure was 1.25 MPa, after reaction for 2.5 h, after cooling and pressure reduction and liquid-solid separation, oxygen leaching residue 292.07 g (Zn 5.5%, Fe 13.2%, Pb 6.20%, S 49.86%) and oxygen leaching solution 2.5 L (H2SO4 50 g / L, Zn 151.56 g / L, Fe 5.6 g / L, Fe 2+ 2.1 g / L) were obtained; the steam consumption of the oxygen pressure leaching process was 81 kg of steam / t of zinc concentrate.

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

[0164] Among them, the content of related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0165] (2) In the 292.07 g of oxygen leaching residue, 1.1 L of waste electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 95 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 58 g / L, Zn 55.12 g / L, Fe 29.09 g / L, Fe 2+ 0.7 g / L) and hot acid residue (172.32 g, Zn 1.46%, Fe 4.7%, Pb 10.51%, S 82.7%); the hot acid residue was sent to a flotation process to recover lead.

[0166] (3) To the 1.1 L of hot acid solution, 47.35 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (23.67 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 73.52 g / L, Fe 29.67 g / L, Fe 2+ 0.66 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

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

[0168] (5) The reduction residue (107.4 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0169] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa. After cooling and pressure reduction, a deironing liquid (1.1 L, Zn 81.11 g / L, Fe 1.51 g / L, H2SO4 39 g / L) and an iron residue (57.11 g, Zn 0.19%, Fe 56.57%, S 1.15%) were obtained; wherein the steam consumption of the deironing process was 1.09 t of steam / t of zinc concentrate.

[0170] (6) The iron residue was sent to subsequent resource recovery;

[0171] The deironing liquid was returned to the oxygen pressure leaching.

[0172] It was calculated that in the above entire process, the recovery rate of Zn was 98.97%, and the recovery rate of Fe was 75.87%.

[0173] Comparative Example 3

[0174] The zinc oxygen pressure leaching and deironing method of the present comparative example comprises the following steps:

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

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

[0177] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0178] (2) In the 292.14 g of oxygen leaching residue, 1.1 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.69 g / L, Fe 28.75 g / L, Fe 2+ 0.7 g / L) and hot acid residue (175.28 g, Zn 1.5%, Fe 4.8%, Pb 10.32%, S 82.4%); the hot acid residue was sent to a flotation process to recover lead.

[0179] (3) To the 1.1 L of hot acid solution, 44.08 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH at the end point was controlled at 1, and after liquid-solid separation, neutralization residue (22.04 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 70.68 g / L, Fe 29.32 g / L, Fe 2+ 0.64 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0180] (4) To the 1.1 L neutralization liquid, 124 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (109.12 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.1 L, Zn 78.41 g / L, Fe 29.91 g / L, Fe 2+ 29.61 g / L).

[0181] (5) The reduction residue (109.12 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0182] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min with the introduction of oxygen with a purity of 99 vol%, the total pressure was controlled at 1.8 MPa, and after cooling and pressure reduction, a deironing liquid (1.1 L, Zn 78.39 g / L, Fe 1.52 g / L, H2SO4 42 g / L) and an iron residue (56.43 g, Zn 0.19%, Fe 56.52%, S 1.23%) were obtained; wherein the steam consumption of the deironing process was 1.09 t of steam / t of zinc concentrate.

[0183] (6) The iron residue was sent to subsequent resource recovery;

[0184] The deironing liquid was returned to the oxygen pressure leaching.

[0185] It was calculated that in the above entire process, the recovery rate of Zn was 98.93%, and the recovery rate of Fe was 75.46%.

[0186] It was found that reducing the pH control end point of the neutralization process caused the generation of a rotten egg smell in the reduction process, increased the amount of zinc concentrate used in the reduction process, and had no obvious effect on the improvement of zinc leaching rate and iron resource utilization rate, and the amount of reduction residue returned increased, which was not conducive to reducing the energy consumption of material transmission.

[0187] Comparative Example 4

[0188] The zinc oxygen pressure leaching and iron removal method of the present comparative example comprises the following steps:

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

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

[0191] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0192] (2) In the 291.93 g of oxygen leaching residue, 1.1 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 92 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 60 g / L, Zn 54.68 g / L, Fe 28.73 g / L, Fe 2+ 0.69 g / L) and hot acid residue (175.16 g, Zn 1.5%, Fe 4.8%, Pb 10.36%, S 82.5%); the hot acid residue was sent to a flotation process to recover lead.

[0193] (3) To the 1.1 L of hot acid solution, 53.88 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH at the end point was controlled at 5, and after liquid-solid separation, neutralization residue (26.94 g, Zn 53%, Fe 12.5%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 74.22 g / L, Fe 25.28 g / L, Fe 2+ 0.61 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0194] (4) To the 1.1 L neutralized liquid, 118 g of zinc concentrate was added, and after reaction at 75 °C for 300 min, liquid-solid separation was performed to obtain a reduction residue (103.84 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) and a reduction liquid (1.1 L, Zn 81.58 g / L, Fe 25.78 g / L, Fe 2+ 25.53 g / L).

[0195] (5) The reduction residue (103.84 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0196] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min with the introduction of oxygen with a purity of 99 vol%, the total pressure was controlled at 1.8 MPa, and after cooling and pressure reduction, a de-ironing liquid (1.1 L, Zn 81.56 g / L, Fe 1.51 g / L, H2SO4 39 g / L) and an iron residue (48.65 g, Zn 0.19%, Fe 56.53%, S 1.2%) were obtained; wherein the steam consumption of the de-ironing process was 1.08 t of steam / t of zinc concentrate.

[0197] (6) The iron residue was sent to subsequent resource recovery;

[0198] The de-ironing liquid was returned to the oxygen pressure leaching.

[0199] It was calculated that in the above entire process, the recovery rate of Zn was 98.93%, and the recovery rate of Fe was 72.23%.

[0200] It was found that during the neutralization process, flocculent iron hydroxide precipitates were formed, which caused the dispersion of iron ions in the open circuit to the hot acid liquid, thereby reducing the utilization rate of iron resources. In addition, due to the loss of iron caused by the neutralization process, the amount of concentrate used for subsequent zinc concentrate reduction was further reduced, that is, the amount of reduction residue produced was reduced, that is, the amount of reduction residue used for oxygen pressure leaching was reduced, thereby increasing the steam consumption of the oxygen pressure leaching process.

[0201] Comparative Example 5

[0202] The zinc oxygen pressure leaching and iron removal method of the present comparative example comprises the following steps:

[0203] (1) 400 g of zinc concentrate (zinc sulfide concentrate), 25.71 g of neutralized residue (Zn 53%, Fe 9%, S 1.5%, Pb 5.05%), and 95.04 g of reduction residue (Zn 42%, Fe 7.5%, S 32%, Pb 3.68%, specific surface area 4.67 m 2 / cm 3), lignin 2.2 g, iron-removing solution 1.1 L (H2SO4 39 g / L, Zn 77.44 g / L, Fe 1.5 g / L), waste electrolyte solution 1.4 L (H2SO4 160 g / L, Zn 43 g / L) were mixed, at this time, the initial sulfuric acid concentration of the reaction system was 106.76 g / L, pure oxygen with a purity of 99 vol% was introduced for oxygen pressure leaching, the oxygen pressure leaching reaction temperature was controlled at 150 ℃, the total pressure was 1.25 MPa, after reaction for 2.5 h, after temperature and pressure reduction and liquid-solid separation, oxygen leaching residue 286.41 g (Zn 5.5%, Fe 13.2%, Pb 6.20%, S 49.86%) and oxygen leaching solution 2.5 L (H2SO4 50 g / L, Zn 148.57 g / L, Fe 5.4 g / L, Fe 2+ 0.7 g / L) were obtained; the steam consumption of the oxygen pressure leaching process was 95 kg of steam / t of zinc concentrate.

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

[0205] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0206] (2) In the 286.41 g of oxygen leaching residue, 1.1 L of waste electrolyte solution (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 85 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 63 g / L, Zn 50.73 g / L, Fe 24.75 g / L, Fe 2+ 0.7 g / L) and hot acid residue (186.17 g, Zn 3.82%, Fe 5.7%, Pb 9.54%, S 81.2%); the hot acid residue was sent to a flotation process to recover lead.

[0207] (3) To the 1.1 L of hot acid solution, 51.43 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (25.71 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 70.72 g / L, Fe 25.24 g / L, Fe 2+ 0.63 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

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

[0209] (5) The reduction residue (95.04 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0210] The reduction liquid was added to an autoclave, and after reaction at 190 °C for 300 min, pure oxygen with a purity of 99 vol% was introduced, and the total pressure was controlled at 1.8 MPa. After cooling and pressure reduction, a deironing liquid (1.1 L, Zn 77.44 g / L, Fe 1.5 g / L, H2SO4 39 g / L) and an iron residue (48.58 g, Zn 0.22%, Fe 56.53%, S 1.2%) were obtained; wherein the steam consumption of the deironing process was 1.08 t of steam / t of zinc concentrate.

[0211] (6) The iron residue was sent to subsequent resource recovery;

[0212] The deironing liquid was returned to the oxygen pressure leaching.

[0213] It was calculated that in the above entire process, the recovery rate of Zn was 97.10%, and the recovery rate of Fe was 73.22%.

[0214] The results showed that a decrease in the hot acid temperature would result in a decrease in the recovery rates of Zn and Fe.

[0215] Comparative Example 6

[0216] The method of zinc oxygen pressure leaching and deironing of the present comparative example comprises the following steps:

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

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

[0219] The content of the related elements in the zinc concentrate is as follows: Zn 49%, Fe 8%, S 29%, Pb 3.24%; the proportion of solid phase with a particle size of-37 μm in the zinc concentrate is greater than 97%, and the specific surface area of the zinc concentrate is 3.98 m 2 / cm 3 .

[0220] (2) In the 291.81 g of oxygen leaching residue, 1.1 L of waste electrolyte (H2SO4 160 g / L, Zn 43 g / L) was added, and after reaction at 97 ℃ for 2 h, solid-liquid separation was performed to obtain hot acid solution (1.1 L, H2SO4 59 g / L, Zn 54.96 g / L, Fe 28.71 g / L, Fe 2+ 0.7 g / L) and hot acid residue (175.09 g, Zn 1.47%, Fe 4.7%, Pb 10.34%, S 82.6%); the hot acid residue was sent to a flotation process to recover lead.

[0221] (3) To the 1.1 L of hot acid solution, 48.16 g of zinc calcine was slowly added at 85 ℃ for 75 min, and the pH was controlled at 2.5 at the end point, and after liquid-solid separation, neutralization residue (24.08 g, Zn 53%, Fe 9%, Pb 5.05%, S 1.5%) and neutralization solution (1.1 L, Zn 73.68 g / L, Fe 29.29 g / L, Fe 2+ 0.64 g / L) were obtained, and the neutralization residue was returned to the oxygen pressure leaching process.

[0222] (4) 121 g of zinc concentrate was added to the 1.1 L neutralization solution, and after reaction at 75°C for 300 min, liquid-solid separation was performed to obtain a reduction residue (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+ 29.57 g / L).

[0223] (5) The reduction residue (106.48 g, Zn 42%, Fe 7.5%, Pb 3.68%, S 32%) was returned to the oxygen pressure leaching;

[0224] The reduction solution was added to an autoclave, and after reaction at 190°C for 300 min with the introduction of oxygen with a purity of 99 vol% and control of the total pressure at 1.8 MPa, cooling and pressure reduction were performed to obtain an iron removal solution (1.1 L, Zn 81.21 g / L, Fe 1.51 g / L, H2SO4 40 g / L) and an iron residue (56.37 g, Zn 0.19%, Fe 56.54%, S 1.21%); wherein the steam consumption of the iron removal process was 1.08 t of steam / t of zinc concentrate.

[0225] (6) The iron residue was sent to subsequent resource recovery;

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

[0227] It was calculated that in the above entire process, the recovery rate of Zn was 98.95%, and the recovery rate of Fe was 75.62%.

[0228] The results show that an increase in the hot acid temperature will increase energy consumption, and the Zn and Fe recovery rates are not obviously improved.

[0229] The content illustrated in the above examples should be understood as the examples being only used to more clearly illustrate the present application, and should not be used to limit the scope of the present application, and after reading the present application, various equivalent modifications of the present application by those skilled in the art all fall within the scope defined by the claims attached to the present application.

Claims

1. A process for zinc oxygen pressure leaching zinc to extract iron, characterized by, The method comprises the following steps: S1, after oxygen pressure leaching of the zinc concentrate to be treated, solid-liquid separation is performed to obtain oxygen leaching residue and oxygen leaching liquid; The zinc concentrate is a zinc sulfide concentrate. S2, after the oxygen smelting slag is subjected to hot acid leaching at 90-95℃, solid-liquid separation is performed, and hot acid leaching residue and hot acid liquid rich in Fe 3+ are obtained; S3, after neutralization treatment of the hot acid liquid, solid-liquid separation is performed to obtain neutralization residue and neutralization liquid with a pH value of 2-4; S4, the neutralization residue is returned to the oxygen pressure leaching procedure in S1; After reduction treatment of the neutralization liquid by a reducing agent, solid-liquid separation is performed to obtain reduction liquid and reduction residue; The reducing agent is ZnS and / or zinc concentrate. S5, after removal of iron from the reduction liquid by the hematite method, solid-liquid separation is performed to obtain iron removal liquid and iron residue; The reduction residue is returned to the oxygen pressure leaching procedure in S1. S6, the iron removal liquid is returned to the oxygen pressure leaching procedure in S1.

2. The method of claim 1, wherein, In S1, when the oxygen pressure leaching is performed, oxygen gas with an oxygen concentration of ≥50 vol% is introduced, a first solution containing sulfuric acid is added, the initial acid concentration of the reaction system is controlled to be 90-125 g / L, the reaction temperature is 135-158 ℃, the total pressure is 1.15-1.35 MPa, and the reaction time is 2-3.5 h.

3. The method of claim 2, wherein, When the oxygen pressure leaching is performed, a surfactant is added to the reaction system, wherein the addition amount of the surfactant is 3-6 wt‰ of the zinc concentrate.

4. The method of claim 3, wherein, The surfactant is lignin.

5. The method of claim 2, wherein, The concentration of sulfuric acid in the first solution is 140-180 g / L.

6. The method of claim 5, wherein, The first solution is waste electrolyte.

7. The method according to any one of claims 2-6, characterized in that, In S2, when the hot acid leaching is performed, the leaching temperature is controlled to be 91-94 ℃.

8. The method of claim 7, wherein, In S2, when the hot acid leaching is performed, the leaching temperature is controlled to be 92-93 ℃.

9. The method according to any one of claims 2-6, characterized in that, The oxygen leaching residue is mixed with a second solution containing sulfuric acid to perform hot acid leaching.

10. The method of claim 9, wherein, The concentration of sulfuric acid 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.

11. The method of claim 10, wherein, The solid-liquid ratio of the oxygen leaching residue to the second solution is 1 g:3-5 mL.

12. The method of claim 9, wherein, The second solution is waste electrolyte.

13. The method of claim 9, wherein, The addition amount of the second solution in S2 is 38-51 vol% of the total addition amount of the first solution in S1 and the second solution in S2.

14. The method of claim 13, wherein, The addition amount of the second solution in S2 is 40-50 vol% of the total addition amount of the first solution in S1 and the second solution in S2.

15. The method of claim 14, wherein, The addition amount of the second solution in S2 is 42-48 vol% of the total addition amount of the first solution in S1 and the second solution in S2.

16. The method of claim 1, wherein, In S2, when the hot acid leaching is performed, the initial acid concentration of the reaction system is 140-180 g / L.

17. The method of claim 1, wherein, In S3, zinc oxide and / or zinc calcine are used to perform the neutralization treatment of the hot acid liquid.

18. The method of claim 17, wherein, When the neutralization treatment is performed, the reaction temperature is controlled to be 80-90 ℃.

19. The method of claim 17, wherein, When the neutralization treatment is performed, the reaction time is controlled to be 60-120 min.

20. The method of claim 1, wherein, In S4, the amount of the reducing agent added at the time of reduction treatment is such that the Fe 3+ is completely converted into Fe 2+ 1.5 to 2.5 times the amount of the reducing agent required.

21. The method of claim 20, wherein, When reduction treatment is performed, the amount of the reducing agent to be added is such that the Fe 3+ is completely converted into Fe 2+ 1.75 to 2.25 times the amount of the reducing agent required.

22. The method of claim 20, wherein, When the reduction treatment is performed, the reaction temperature is controlled to be 70-80 ℃, and the reaction time is 4-6 h.

23. The method of claim 1, wherein, In S5, when the hematite method is used to remove iron, oxygen gas with an oxygen concentration of ≥50 vol% is introduced into the reduction liquid, the reaction temperature is controlled to be 180-200 ℃, the total pressure is 1.6-2.0 MPa, and the reaction time is 4-6 h.

24. The method of any one of claims 1-6, 8, 10-23, wherein, In the zinc concentrate, the content of Zn is 40-60 wt%, the content of Fe is 4-12 wt%, and the content of S is 20-40 wt%.

25. The method of claim 24, wherein, The zinc concentrate contains 48-51wt% of Zn, 6-9wt% of Fe and 28-32wt% of S.

26. The method of any one of claims 1-6, 8, 10-23, wherein, The zinc concentrate contains more than 95wt% of solid phase with particle size of-37μm.

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