Method for removing iron by zinc immersion under zinc oxygen pressure

Abstract

The invention relates to a method for removing iron by zinc leaching under oxygen pressure, which comprises the following steps: after carrying out first-stage oxygen pressure leaching on zinc concentrate to be treated, carrying out solid-liquid separation to obtain first leaching residue and first leaching liquid; after second-stage oxygen pressure leaching is conducted on the first leaching residues, solid-liquid separation is conducted, and second leaching residues and second leaching liquid are obtained; returning the secondary leaching liquid to a first-stage oxygen pressure leaching process; carrying out hot acid leaching on the secondary leaching residues, and carrying out solid-liquid separation to obtain hot acid residues and hot acid liquid; neutralizing the hot acid liquid to obtain neutralized slag and neutralized liquid; returning the neutralized slag to a first-stage oxygen pressure leaching process; after the neutralization solution is subjected to reduction treatment, reducing liquid and reducing slag are obtained; after the reduction liquid is subjected to iron removal through a hematite method, solid-liquid separation is conducted, and iron-removed liquid and iron slag are obtained; the reducing slag is returned to the second-stage oxygen pressure leaching procedure; and the iron removal liquid is returned to the first-stage oxygen pressure leaching process. According to the method, loss of valuable elements such as zinc can be reduced while open circuit of iron in the zinc oxygen pressure leaching system is achieved, and the recovery rate of Zn can reach 99% or above.

Description

Technical Field

[0001] The invention relates to a method for removing iron from zinc 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 the problem of 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. Therefore, there is an urgent need to develop more advanced zinc oxygen pressure leaching process technology, which can reduce the volume of iron removal and improve the iron removal efficiency while achieving efficient zinc leaching, realize efficient leaching of other valuable metals such as copper, take into account the self-heating reaction of pressure leaching, and improve the economic benefits of the zinc smelting system.

[0003] 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

[0004] In view of the shortcomings of the prior art, the object of the present invention is to provide a method for removing iron from zinc by zinc oxygen pressure leaching, so as to reduce the loss of valuable elements such as zinc while removing iron from the system by open circuit zinc oxygen pressure leaching.

[0005] In order to solve the above technical problems, the technical solution of the present invention is as follows:

[0006] A method for removing iron from zinc by zinc-oxygen pressure leaching, comprising the following steps:

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

[0008] Wherein, the zinc concentrate is zinc sulfide concentrate;

[0009] S2, subjecting the first leaching residue to two-stage oxygen pressure leaching, and then separating the solid and liquid to obtain a second leaching residue and a second leaching liquid;

[0010] S3, returning the secondary leaching solution to the first oxygen pressure leaching process in S1;

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

[0012] S4, after neutralizing the hot acid liquid, separate the solid and liquid to obtain a neutralized slag and a neutralized liquid with a pH value of 2 to 4;

[0013] S5, returning the neutralized slag to the oxygen pressure leaching process in S1;

[0014] After the neutralized liquid is reduced by a reducing agent, solid-liquid separation is performed to obtain a reduced liquid and reduced slag;

[0015] Wherein, the reducing agent is ZnS and / or zinc concentrate;

[0016] S6, 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;

[0017] Returning the reduced slag to the second-stage oxygen pressure leaching process in S2;

[0018] S7, returning the iron removal solution to the oxygen pressure leaching process in S1.

[0019] Thus, the zinc concentrate is mixed with the returned secondary leaching solution, neutralization slag, iron removal liquid and other materials, and after a first oxygen pressure leaching treatment, a first leaching slag and a first leaching solution are obtained; the first leaching solution can be sent to the downstream side for pre-neutralization, purification, electrolysis, casting and other processes, and the first leaching slag is further mixed with the reducing slag, the first acid solution and other materials, and after a second oxygen pressure leaching treatment, a second leaching solution and a second leaching slag are obtained. Subsequently, after the valuable elements such as iron and zinc in the secondary leaching slag are fully leached by hot acid leaching, the hot acid liquid is neutralized to prepare for the next reduction treatment, 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 to allow sulfur to exist in the reducing slag in the form of elemental 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 open circuit of iron in the zinc oxygen pressure leaching system; the iron removal liquid returns to the first oxygen pressure leaching process in S1, and the acid therein can participate in the first oxygen pressure leaching, and the zinc and other valuable metals therein can be further enriched in the first leaching liquid and enter the subsequent process, effectively reducing the loss of zinc and other valuable metals. At the same time, the reduction slag returns to the second oxygen pressure leaching process in S2. 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 and activity, the second oxygen pressure leaching process reaction (1) and reaction (6) are easier to proceed, and the heat is released stably to maintain the progress of the second oxygen pressure leaching reaction, which helps to reduce the external heat required for the second oxygen pressure leaching process. For example, the steam consumption can be reduced from 0.1t steam / t-zinc concentrate to 0.01t steam / t-zinc concentrate.

[0020] Furthermore, during a stage of oxygen pressure leaching, an oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced, the initial acid concentration of the reaction system (in terms of H2SO4 concentration) is controlled to be 30-60 g / L, preferably 35-55 g / L, the reaction temperature is 140-160°C, preferably 145-155°C, the total pressure is 1.0-1.3 MPa, preferably 1.1-1.2 MPa, and the reaction time is 1-2 h, preferably 1.2-1.8 h;

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

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

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

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

[0025] Preferably, the concentration of sulfuric acid in the first solution is 140-180 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.

[0026] Further, in S3, when hot acid leaching is performed, the leaching temperature is controlled to be 80 to 99° C., preferably 90 to 95° C.;

[0027] Preferably, the secondary leaching residue 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, and the solid-liquid ratio of the secondary leaching residue to the second solution is 1 g:2-5 mL; further preferably, the second solution is a waste electrolyte.

[0028] Preferably, the amount of the second solution added in S3 is 35-45 vol% of the total amount of the first solution in S2 and the second solution in S3, and more preferably 38-42 vol%. Thus, while satisfying oxygen pressure leaching and hot acid leaching, the volume of the solution to be deironed can be greatly reduced, the heat consumption required for deironing can be reduced, and the deironing cost can be significantly reduced in the iron source opening process.

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

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

[0031] Preferably, during the neutralization treatment, the reaction temperature is controlled to be 80-90°C;

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

[0033] Furthermore, in S5, when the reduction treatment is performed, the amount of the reducing agent added is such that Fe 3+ Completely converted to Fe 2+ 1.2-1.8 times, preferably 1.35-1.65 times, the amount of the required reducing agent;

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

[0035] Furthermore, in S6, when the hematite method is used for iron removal, an oxygen-rich gas with an oxygen concentration ≥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, more preferably 4.5-5.5 h.

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

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

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

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

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

[0041] ZnO+H2SO4=ZnSO4+H2O (4)

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

[0043] S+H2O+O2=H2SO4 (6)

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

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

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

[0047] Furthermore, in the zinc concentrate, the content of solid matter with a particle size of -44 μm is greater than 97 wt%.

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

[0049] Compared with the prior art, the present invention has the following beneficial effects:

[0050] (1) The method of the present invention can realize iron circuit opening 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 98.5%.

[0051] (2) The present invention focuses on the hot acid leaching of iron in the secondary leaching residue. The volume of hot acid solution produced by hot acid leaching is small, which greatly reduces the volume of subsequent iron removal solution. The heat required for the subsequent hematite iron removal process is greatly reduced, thereby significantly reducing the cost of iron removal. 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 ≤0.2%, and the S content is ≤1.2%. The loss of valuable elements such as zinc and sulfur in the zinc smelting system is small.

[0052] (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 second-stage oxygen pressure leaching process, which can make the second-stage oxygen pressure leaching reaction easier to carry out, and help to 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

[0053] Figure 1 The present invention discloses a flow chart of a process for removing iron by zinc-oxygen pressure leaching. DETAILED DESCRIPTION

[0054] 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.

[0055] Example 1

[0056] The method for removing iron from zinc by zinc oxygen pressure leaching of the present embodiment comprises the following steps:

[0057] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 )400g, neutralized slag (16.33g, Zn 53%, Fe 9%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (1L, H2SO4 40g / L, Zn 72.6g / L, Fe 1.5g / L, Cu 0.25g / L), second immersion solution (1.4L, H2SO4 42g / L, Zn99.95g / L, Fe 10.28g / L, Cu 0.65g / L), oxygen-enriched gas with an oxygen concentration of 99vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150°C, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (275g, Zn10.85%, Fe22.65%, Cu0.3%, S40.88%) and a leaching solution (2.4L, H2SO4 12.5g / L, Zn 161.4g / L, Fe (total iron) 1.6g / L, Fe 2+ 0.52g / L, Cu 1.08g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0058] (2) All the above-mentioned leaching slag and reduction slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68 m 2 / cm 3 ) was mixed, lignin (1.6 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 1.78, and oxygen-rich gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching. The reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount. The reaction time was 2 h. After cooling and pressure reduction and liquid-solid separation, the second leaching residue (290.7 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching liquid (1.4 L, H2SO4 42 g / L, Zn 99.95 g / L, Fe 10.28 g / L, Fe 2+ 1.1g / L, Cu 0.65g / L), wherein the steam consumption in the second stage oxygen pressure leaching process is 0.01t steam / t zinc concentrate.

[0059] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0060] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (290.7g, Zn1%, Fe 21.01%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 45.33g / L, Fe 50.09g / L, Fe 2+ 0.8g / L, Cu 0.17g / L) and hot acid slag (116.3g, Zn0.5%, Fe 5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0061] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.33g / L, Fe 50.09g / L, Fe 2+ 0.8g / L, Cu0.17g / L) was slowly added with 32.65g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85℃, and the reaction was carried out for 75min. The end point pH was controlled to be 2.5. Neutralized slag (16.33g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and neutralized solution (1L, Zn 63.01g / L, Fe 51.09g / L, Fe 2+ 0.7g / L, Cu 0.26g / L).

[0062] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0063] The above neutralization solution (1L, Zn 63.01g / L, Fe 51.09g / L, Fe 2+ 0.7g / L, Cu 0.26g / L) for reduction treatment, wherein 140g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (128.8g, Zn42%, Fe 7.5%, Cu 0.53%, S32%) and reduction solution (1L, Zn72.60g / L, Fe53.15g / L, Fe 2+ 53.02g / L, Cu 0.25g / L).

[0064] (6) returning the reduced slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the second-stage oxygen pressure leaching process;

[0065] The reduced solution (1L, Zn 72.61g / L, Fe 53.15g / L, Fe 2+(53.02g / L, Cu 0.26g / L) was added into the autoclave, and oxygen with a purity of 99vol% was introduced at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 72.60g / L, Fe1.5g / L, Cu 0.25g / L, H2SO4 40g / L) and iron slag (90.93g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. Among them, the steam consumption of the iron removal process was 1.0t steam / t zinc concentrate. (7) The above iron removal liquid was returned to the first oxygen pressure leaching process.

[0066] It is calculated that in the above whole process, the recovery rate of Zn is 99.7%; the recovery rate of iron is 79.0%.

[0067] Comparative Example 1

[0068] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0069] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 )400g, lignin (1.6g), iron-removing liquid (1L, H2SO4 70g / L, Zn65.39g / L, Fe4.5g / L, Cu 0.23g / L), second leaching liquid (1.4L, H2SO4 63g / L, Zn110.04g / L, Fe 9.72g / L, Cu 0.76g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for one-stage oxygen pressure leaching, the reaction temperature was 150℃, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (272g, Zn9.85%, Fe11.65%, Cu0.3%, S40.88%) and a leaching liquid (2.4L, H2SO4 46.8g / L, Zn 162.17g / L,Fe 8.88g / L,Fe 2+ 4.85g / L, Cu 1.11g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0070] (2) All the above-mentioned leaching slag and reduction slag (180 g, Zn 40%, Fe 7.6%, Cu 0.53%, S 30%, specific surface area 4.65 m 2 / cm 3) was mixed, lignin (1.6 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 1.51, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount, and the reaction time was 2 h. After cooling and pressure reduction and liquid-solid separation, the second leaching residue (352.56 g, Zn 1.2%, Fe 10.01%, Cu 0.1%, S 51.61%) and the second leaching liquid (1.4 L, H2SO4 63 g / L, Zn 110.04 g / L, Fe 9.72 g / L, Fe 2+ 0.9g / L, Cu 0.76g / L), the steam consumption of the second stage oxygen pressure leaching is 0.015t steam / t zinc concentrate.

[0071] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0072] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (352.56g, Zn 1.2%, Fe 10.01%, Cu 0.1%, S 51.61%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 46.38g / L, Fe 31.76g / L, Fe 2+ 0.8g / L, Cu 0.21g / L) and hot acid slag (141.02g, Zn0.5%, Fe5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0073] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 46.38g / L, Fe 31.76g / L, Fe 2+ 0.8g / L, Cu0.21g / L) was reduced with zinc concentrate, wherein 200g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduced slag (180g, Zn40%, Fe7.6%, Cu0.53%, S30%) and reduced solution (1L, Zn65.98g / L, Fe32.40g / L, Fe 2+ 32.30 g / L, Cu 0.23 g / L).

[0074] (5) returning the reduced slag (180 g, Zn 40%, Fe 7.6%, Cu 0.53%, S 30%) to the second-stage oxygen pressure leaching process;

[0075] The above reduced solution (1L, Zn 65.98g / L, Fe32.40g / L, Fe 2+ 32.30g / L, Cu 0.23g / L) was added into the autoclave, and the oxygen concentration was 99vol% at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 65.95g / L, Fe4.5g / L, Cu 0.23g / L, H2SO4 70g / L) and iron slag (59.29g, Zn0.2%, Fe 53.25%, S1.2%), and the iron slag was sent to subsequent resource recovery. Among them, the steam consumption of the iron removal process is 1.0t steam / t zinc concentrate.

[0076] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0077] According to calculation, in the above whole process, the recovery rate of Zn is 99.7%; the recovery rate of iron is 40.9%.

[0078] The results showed that the pre-neutralization process was abandoned and the hot acid solution was directly reduced. A strong smell of rotten eggs was generated during the reduction process, which increased the amount of zinc concentrate and worsened the operating environment. Increasing the amount of zinc concentrate used for reduction will also increase the initial acidity of hematite before iron removal, which has little effect on the zinc leaching rate, but greatly increases the iron content of the first-stage leaching solution, increases the pressure of subsequent iron removal and purification of the first-stage leaching solution, and significantly reduces the overall iron resource recovery rate to 40.9%. The amount of steam consumed in the second-stage oxygen pressure leaching also increased.

[0079] Comparative Example 2

[0080] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0081] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3)400g, neutralized slag (15.51g, Zn 53%, Fe 9%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (1L, H2SO4 40g / L, Zn 72.45g / L, Fe 1.5g / L, Cu 0.25g / L), second immersion solution (1.4L, H2SO4 42g / L, Zn102.58g / L, Fe 10.38g / L, Cu 0.67g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150℃, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (275g, Zn10.85%, Fe22.65%, Cu0.3%, S40.88%) and a leaching solution (2.4L, H2SO4 12.5g / L, Zn 162.68g / L, Fe1.6g / L, Fe 2+ 0.52g / L, Cu 1.09g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0082] (2) All the above-mentioned leaching slag and reduction slag (138g, Zn42%, Fe7.5%, Cu0.53%, S32%, specific surface area4.66m 2 / cm 3 ) was mixed, lignin (1.6 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 1.70, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount, and the reaction time was 2 h. After cooling, pressure reduction and liquid-solid separation, the second leaching residue (297.36 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching liquid (1.4 L, H2SO4 42 g / L, Zn 102.58 g / L, Fe 10.38 g / L, Fe 2+ 1.1g / L, Cu 0.67g / L), the steam consumption in the second-stage oxygen pressure leaching process is 0.011t steam / t zinc concentrate.

[0083] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0084] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (297.36g, Zn1%, Fe 21.01%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 45.38g / L, Fe 51.23g / L, Fe2+ 0.8g / L, Cu 0.18g / L) and hot acid slag (118.94g, Zn0.5%, Fe 5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0085] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.38g / L, Fe 51.23g / L, Fe 2+ 0.8g / L, Cu0.18g / L) was slowly added with 31.02g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85℃, and the reaction was carried out for 75min. The pH value was controlled at 1 at the end point. The neutralized slag (15.51g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and the neutralized solution (1L, Zn 62.18g / L, Fe 51.74g / L, Fe 2+ 0.7g / L, Cu 0.26g / L).

[0086] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0087] The above neutralization solution (1L, Zn 62.18g / L, Fe 51.74g / L, Fe 2+ 0.7g / L, Cu 0.26g / L) for reduction treatment, wherein 150g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (138g, Zn42%, Fe 7.5%, Cu 0.53%, S32%) and reduction solution (1L, Zn72.47g / L, Fe53.15g / L, Fe 2+ 53.02g / L, Cu 0.26g / L).

[0088] (6) returning the reduced slag (138 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the second-stage oxygen pressure leaching process;

[0089] The reduced solution (1L, Zn 72.47g / L, Fe 53.15g / L, Fe 2+53.02g / L, Cu 0.26g / L) was added into the autoclave, and the oxygen concentration was 99vol% at 190℃, and the total pressure was controlled at 1.8MPa. After 300min of reaction, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 72.45g / L, Fe1.5g / L, Cu 0.25g / L, H2SO4 40g / L) and iron slag (90.93g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. The steam consumption of the iron removal process is 1.0t steam / t zinc concentrate.

[0090] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0091] It is calculated that in the above whole process, the recovery rate of Zn is 99.7%; the recovery rate of iron is 79.0%.

[0092] 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.

[0093] Comparative Example 3

[0094] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0095] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 ) 400g, neutralized slag (19.59g, Zn 53%, Fe 12.5%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (1L, H2SO4 40g / L, Zn 73.88g / L, Fe 1.5g / L, Cu 0.25g / L), second immersion solution (1.4L, H2SO4 42g / L, Zn92.16g / L, Fe 10.01g / L, Cu 0.58g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150℃, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (276g, Zn10.85%, Fe22.65%, Cu0.3%, S40.88%) and a leaching solution (2.4L, H2SO4 12.5g / L, Zn 158.06g / L, Fe1.6g / L, Fe 2+ 0.52g / L, Cu 1.05g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0096] (2) All the above-mentioned leaching slag and reduction slag (101.2 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68 m 2 / cm 3 ) was mixed, lignin (1.6 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled at 2.06, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount, and the reaction time was 2 h. After cooling, pressure reduction and liquid-solid separation, the second leaching residue (271.58 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching liquid (1.4 L, H2SO4 42 g / L, Zn 92.16 g / L, Fe 10.01 g / L, Fe 2+ 1.1g / L, Cu 0.58g / L), wherein the steam consumption in the second-stage oxygen pressure leaching process is 0.02t steam / t zinc concentrate.

[0097] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0098] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (271.58g, Zn1%, Fe 21.01%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 45.17g / L, Fe 46.79g / L, Fe 2+ 0.8g / L, Cu 0.16g / L) and hot acid slag (108.63g, Zn0.5%, Fe 5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0099] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 45.17g / L, Fe 46.79g / L, Fe 2+ 0.8g / L, Cu0.16g / L) was slowly added with zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) 39.18g at 85℃, reacted for 75min, pH=5 was controlled at the end point, and neutralized slag (19.59g, Zn 53%, Fe 12.5%, Cu 0.57%, S1.5%) and neutralized solution (1L, Zn 66.39g / L, Fe42.95g / L, Fe 2+ 0.7g / L, Cu 0.26g / L).

[0100] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0101] The above neutralization solution (1L, Zn 66.39g / L, Fe42.95g / L, Fe 2+ 0.7g / L, Cu 0.26g / L) was used for reduction treatment, wherein 110g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, the reduction slag (101.20g, Zn42%, Fe 7.5%, Cu 0.53%, S32%, specific surface area 4.58m 2 / cm 3 ) and reducing solution (1L, Zn 73.94g / L, Fe43.81g / L, Fe 2+ 43.75g / L, Cu0.26g / L).

[0102] (6) returning the reduced slag (101.20 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the second-stage oxygen pressure leaching process;

[0103] The reduced solution (1L, Zn 73.94g / L, Fe 43.81g / L, Fe 2+ 43.75g / L, Cu 0.26g / L) was added into the autoclave, and the oxygen concentration was 99vol% at 190℃, and the total pressure was controlled at 1.8MPa. After 300min of reaction, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 73.88g / L, Fe1.5g / L, Cu 0.25g / L, H2SO4 40g / L) and iron slag (74.95g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. The steam consumption of the iron removal process is 1.0t steam / t zinc concentrate.

[0104] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0105] It is calculated that in the above whole process, the recovery rate of Zn is 99.7%; the recovery rate of iron is 71.1%.

[0106] The results show that flocculent iron hydroxide precipitates in the neutralization process of step (4), which causes the iron ions in the hot acid solution to disperse, thereby reducing the utilization rate of iron resources. In addition, due to the iron loss caused by the neutralization process, the amount of concentrate used for subsequent zinc concentrate reduction is further reduced, that is, the amount of reduction slag produced is reduced, that is, the amount of reduction slag returned to the second stage oxygen pressure leaching is reduced, resulting in an increase in the amount of steam used in the second stage oxygen pressure leaching process.

[0107] Comparative Example 4

[0108] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0109] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 )400g, lignin (1.6g), iron-removing liquid (1L, H2SO4 40g / L, Zn79.99g / L, Fe 1.5g / L, Cu 0.25g / L), second leaching liquid (1.4L, H2SO4 42g / L, Zn99.81g / L, Fe 10.21g / L, Cu 0.64g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for one-stage oxygen pressure leaching, the reaction temperature was 150℃, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (273g, Zn10.85%, Fe22.65%, Cu0.3%, S40.88%) and a leaching liquid (2.4L, H2SO4 12.5g / L, Zn 160.88g / L,Fe 1.6g / L,Fe 2+ 0.53g / L, Cu 1.08g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0110] (2) All the above-mentioned leaching slag and reduction slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68 m 2 / cm 3 ), neutralized slag (16.33g, Zn53%, Fe 9%, Cu 0.57%, S1.5%), added lignin (1.6g), added 1.4L spent electrolyte (Zn 43g / L, H2SO4 160g / L), at this time the acid-zinc molar ratio was 1.79, oxygen gas with an oxygen concentration of 99vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150℃, the total pressure was 1.25MPa, the reaction time was 2h, and after cooling and pressure reduction and liquid-solid separation, the second leaching slag (289.30g, Zn4.2%, Fe 21.01%, Cu0.1%, S 51.62%) and the second leaching solution (1.4L, H2SO4 42g / L, Zn 99.81g / L, Fe 10.21g / L, Fe 2+ 1.1g / L, Cu 0.64g / L), the steam consumption in the second stage oxygen pressure leaching process is 0.011t steam / t zinc concentrate.

[0111] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0112] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (289.30g, Zn 4.2%, Fe 21.01%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 52.72g / L, Fe 49.84g / L, Fe 2+ 0.8g / L, Cu 0.17g / L) and hot acid slag (115.72g, Zn2.3%, Fe5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0113] (4) Add the hot acid solution (1L, H2SO4 40g / L, Zn 52.72g / L, Fe 49.84g / L, Fe 2+ 0.8 g / L, Cu0.17 g / L) was slowly added with 32.65 g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85°C, and the reaction was continued for 75 min. The pH value was controlled at 2.5 at the end point. The neutralized slag (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and the neutralized solution (1 L, Zn 70.40 g / L, Fe 50.84 g / L, Fe 2+ 0.7g / L, Cu 0.25g / L).

[0114] (5) returning the neutralized slag to the second stage for oxygen pressure leaching;

[0115] The above neutralization solution (1L, Zn 70.40g / L, Fe 50.84g / L, Fe 2+ 0.7g / L, Cu 0.25g / L) for reduction treatment, wherein 140g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (128.8g, Zn42%, Fe 7.5%, Cu 0.53%, S32%) and reduction solution (1L, Zn80.01g / L, Fe51.25g / L, Fe 2+ 51.15g / L, Cu 0.26g / L).

[0116] (6) returning the reduced slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the second-stage oxygen pressure leaching process;

[0117] The reduced solution (1L, Zn 80.01g / L, Fe 51.25g / L, Fe 2+51.15g / L, Cu 0.26g / L) was added into the autoclave, and the oxygen concentration was 99vol% at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 79.99g / L, Fe1.5g / L, Cu 0.25g / L, H2SO4 40g / L) and iron slag (87.68g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. The steam consumption of the iron removal process is 1.0t steam / t zinc concentrate.

[0118] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0119] According to calculation, in the above whole process, the recovery rate of Zn is 98.9% and the recovery rate of iron is 79.0%. The recovery rate of zinc is reduced.

[0120] Comparative Example 5

[0121] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0122] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 ) 400g, neutralized slag (16.33g, Zn 53%, Fe 9%, S1.5%), reduced slag (128.8g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68m 2 / cm 3 ), add lignin (1.6g), add iron-removing liquid (1L, H2SO4 40g / L, Zn 71.82g / L, Fe 1.5g / L, Cu0.25g / L), second leaching liquid (1.4L, H2SO4 42g / L, Zn64.20g / L, Fe 13.35g / L, Cu 0.35g / L), pass oxygen gas with an oxygen concentration of 99vol% for one-stage oxygen pressure leaching, the reaction temperature is 150℃, the total pressure is 1.25MPa, the reaction time is 1.5h, after cooling and pressure reduction and liquid-solid separation, a leaching residue (275g, Zn11.36%, Fe22.65%, Cu0.3%, S40.88%) and a leaching liquid (2.4L, H2SO4 13.6g / L, Zn 162.17g / L,Fe5.6g / L,Fe 2+ 1.65g / L, Cu 1.11g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0123] (2) Add lignin (1.6 g) and 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) to all the above-mentioned first leaching residues. At this time, the acid-zinc molar ratio is 4.78. Oxygen gas with an oxygen concentration of 99 vol% is introduced for two-stage oxygen pressure leaching. The reaction temperature is controlled to be 150°C, the total pressure is 1.25 MPa, and the reaction time is 2 h. After cooling and pressure reduction and liquid-solid separation, the second leaching residue (198 g, Zn 1.5%, Fe 16.08%, Cu 0.1%, S 51.62%) and the second leaching solution (1.4 L, H2SO4 73 g / L, Zn 64.2 g / L, Fe 21.36 g / L, Fe 2+ 2.1g / L, Cu 0.35g / L), the second stage steam consumption is 0.09t steam / t zinc concentrate.

[0124] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0125] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (198g, Zn 1.5%, Fe 16.08%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40.5g / L, Zn 45.38g / L, Fe 26.11g / L, Fe 2+ 0.85g / L, Cu 0.12g / L) and hot acid slag (79.2g, Zn0.52%, Fe5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0126] (4) Add the above hot acid solution (1L, H2SO4 40.5g / L, Zn 45.38g / L, Fe 26.11g / L, Fe 2+ 0.85g / L, Cu 0.12g / L) was slowly added with 33.06g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85℃, and the reaction was carried out for 75min. The pH value was controlled at 2.5 at the end point. Neutralized slag (16.53g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and neutralized solution (1L, Zn 63.28g / L, Fe 26.03g / L, Fe 2+ 0.75g / L, Cu 0.20g / L).

[0127] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0128] The above neutralization solution (1L, Zn 63.28g / L, Fe26.03g / L, Fe 2+0.75g / L, Cu0.20g / L) was used for reduction treatment, wherein 140g of zinc concentrate was added, the reaction temperature was 85°C, and the reaction time was 300min. After the reaction was completed, liquid-solid separation was performed to obtain reduction slag (128.8g, Zn42%, Fe7.5%, Cu0.53%, S32%) and reduction solution (1L, Zn72.88g / L, Fe27.16g / L, Fe 2+ 27.02g / L, Cu 0.20g / L).

[0129] (6) returning the reduced slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the first oxygen pressure leaching process;

[0130] The reduced solution (1L, Zn 72.88g / L, Fe 27.16g / L, Fe 2+ 27.02g / L, Cu 0.20g / L) was added into the autoclave, at 190℃, the oxygen concentration was 99vol%, the total pressure was controlled at 1.8MPa, after 300min of reaction, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 71.82g / L, Fe1.5g / L, Cu 0.25g / L, H2SO4 40g / L) and iron slag (46.47g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. The steam consumption of iron removal process is 1.0t steam / t zinc concentrate.

[0131] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0132] It was calculated that in the above whole process, the recovery rate of Zn was 99.7%; and the recovery rate of iron was 57.31%.

[0133] Returning the reduced slag produced in step (5) to the first stage oxygen pressure leaching has little effect on the zinc leaching rate, but has a great impact on the steam consumption in the second stage oxygen pressure leaching process, significantly increasing the steam consumption, which is not conducive to energy saving and consumption reduction.

[0134] Comparative Example 6

[0135] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0136] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3)400g, neutralized slag (32.65g, Zn 53%, Fe 9%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (2L, H2SO4 30g / L, Zn 69.4g / L, Fe 1.5g / L, Cu 0.17g / L), second immersion solution (0.4L, H2SO4 52g / L, Zn179.24g / L, Fe 36.20g / L, Cu2.27g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150℃, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (277g, Zn13.26%, Fe22.65%, Cu0.3%, S40.88%) and a leaching solution (2.4L, H2SO4 13.4g / L, Zn 132.36g / L, Fe1.6g / L, Fe 2+ 0.53g / L, Cu 1.15g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0137] (2) All the above-mentioned leaching slag and reduction slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%, specific surface area 4.68 m 2 / cm 3 ) was mixed, lignin (1.6 g) was added, 0.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 0.47, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C, the total pressure was 1.25 MPa, the reaction time was 2 h, and after cooling and pressure reduction and liquid-solid separation, the second leaching residue (292.18 g, Zn 4.5%, Fe 21.01%, Cu 0.1%, S 51.62%) and the second leaching solution (0.4 L, H2SO4 52 g / L, Zn 179.24 g / L, Fe 36.20 g / L, Fe 2+ 1.3g / L, Cu 2.27g / L), wherein the steam consumption in the second-stage oxygen pressure leaching process is 0.01t steam / t zinc concentrate.

[0138] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0139] 2L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (292.18g, Zn 4.5%, Fe 21.01%, Cu 0.1%, S 51.62%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (2L, H2SO4 40g / L, Zn 46.94g / L, Fe 25.17g / L, Fe2+ 0.8g / L, Cu 0.09g / L) and hot acid slag (116.87g, Zn3.5%, Fe5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0140] (4) Add the hot acid solution (2L, H2SO4 40g / L, Zn 46.94g / L, Fe 25.17g / L, Fe 2+ 0.8g / L, Cu0.09g / L) was slowly added with 65.31g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85℃, and the reaction was carried out for 75min. The pH value was controlled at 2.5 at the end point. Neutralized slag (32.65g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and neutralized liquid (2L, Zn 64.63g / L, Fe 25.67g / L, Fe 2+ 0.7g / L, Cu 0.17g / L).

[0141] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0142] The above neutralization solution (2L, Zn 64.63g / L, Fe25.67g / L, Fe 2+ 0.7g / L, Cu 0.17g / L) for reduction treatment, wherein 140g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (128.8g, Zn42%, Fe 7.5%, Cu 0.53%, S32%) and reduction solution (2L, Zn69.43g / L, Fe26.19g / L, Fe 2+ 26.15 g / L, Cu 0.17 g / L).

[0143] (6) returning the reduced slag (128.8 g, Zn 42%, Fe 7.5%, Cu 0.53%, S 32%) to the second-stage oxygen pressure leaching process;

[0144] The reduced solution (2L, Zn 69.43g / L, Fe 26.19g / L, Fe 2+26.15g / L, Cu 0.17g / L) was added into the autoclave, and the oxygen concentration was 99vol% at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (2L, Zn 69.4g / L, Fe1.5g / L, Cu 0.17 / L, H2SO4 30g / L) and iron slag (89.59g, Zn0.2%, Fe 55.65%, S1.2%), and the iron slag was sent to subsequent resource recovery. The steam consumption of the iron removal process is 2.25t steam / t zinc concentrate.

[0145] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0146] It is calculated that in the above whole process, the recovery rate of Zn is 98.4%; the recovery rate of iron is 73.2%.

[0147] The results showed that the volume of the spent electrolyte in step (2) was controlled to 0.4 L, and the volume of the spent electrolyte in the hot acid leaching process in step (3) was controlled to 2 L. The recovery rate of zinc decreased, and the volume of the iron removal liquid doubled, which significantly increased the amount of steam used in the hematite iron removal process.

[0148] Comparative Example 7

[0149] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0150] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 )400g, neutralized slag (16.33g, Zn 53%, Fe 9%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (1L, H2SO4 25g / L, Zn 63.7g / L, Fe 27.5g / L, Cu 0.21g / L), second immersion solution (1.4L, H2SO4 70g / L, Zn65.22g / L, Fe 9.03g / L, Cu 0.39g / L), oxygen gas with an oxygen concentration of 99vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150°C, the total pressure was 1.25MPa, the reaction time was 1.5h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (272g, Zn10.92%, Fe22.72%, Cu0.3%, S40.85%) and a leaching solution (2.4L, H2SO4 12.7g / L, Zn 137.48g / L, Fe (total iron) 6.7g / L, Fe 2+ 4.89g / L, Cu 0.92g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0151] (2) All the above-mentioned leaching slag and reduction slag (18.40 g, Zn 16.5%, Fe 7.5%, Cu 0.53%, S 45.3%, specific surface area 4.56 m 2 / cm 3 ) was mixed, lignin (1.2 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 4.57, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount, and the reaction time was 2 h. After cooling, pressure reduction and liquid-solid separation, the second leaching residue (209.1 g, Zn 1%, Fe 21.01%, Cu 0.1%, S 52.64%) and the second leaching liquid (1.4 L, H2SO4 70 g / L, Zn 65.22 g / L, Fe 9.03 g / L, Fe 2+ 1.2g / L, Cu 0.39g / L), wherein the steam consumption in the second stage oxygen pressure leaching process is 0.06t steam / t zinc concentrate.

[0152] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0153] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (209.1g, Zn1%, Fe 21.01%, Cu 0.1%, S 52.64%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 44.67g / L, Fe 36.02g / L, Fe 2+ 0.9g / L, Cu 0.13g / L) and hot acid slag (83.64g, Zn0.5%, Fe5.1%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0154] (4) Add the above hot acid solution (1L, H2SO4 40g / L, Zn 44.67g / L, Fe 36.02g / L, Fe 2+ 0.9g / L, Cu0.13g / L) was slowly added with 32.65g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85℃, and the reaction was continued for 75min. The end point pH was controlled to be 2.5. The neutralized slag (16.33g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and the neutralized solution (1L, Zn 62.35g / L, Fe 36.74g / L, Fe 2+ 0.7g / L, Cu 0.21g / L).

[0155] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0156] The above neutralization solution (1L, Zn 62.35g / L, Fe36.74g / L, Fe 2+ 0.7g / L, Cu 0.21g / L) for reduction treatment, wherein 20g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (18.40g, Zn16.5%, Fe 7.5%, Cu 0.53%, S45.3%) and reduction solution (1L, Zn63.73g / L, Fe37.48g / L, Fe 2+ 11.24 g / L, Cu 0.21 g / L).

[0157] (6) returning the reduced slag (18.40 g, Zn 16.5%, Fe 7.5%, Cu 0.53%, S 45.3%) to the second-stage oxygen pressure leaching process;

[0158] The reduced solution (1L, Zn 63.73g / L, Fe 37.48g / L, Fe 2+ 11.24g / L, Cu 0.21g / L) was added into the autoclave, and oxygen with a purity of 99vol% was introduced at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 63.7g / L, Fe27.5g / L, Cu 0.21g / L, H2SO4 25g / L) and iron slag (17.85g, Zn0.2%, Fe 56.74%, S1.2%), and the iron slag was sent to subsequent resource recovery. Among them, the steam consumption of the iron removal process is 2.1t steam / t zinc concentrate.

[0159] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0160] It is calculated that in the above whole process, the recovery rate of Zn is 99.7%; the recovery rate of iron is 20.8%.

[0161] The results show that reducing the amount of zinc concentrate added in step (5) to 20 g directly causes the Fe content in the reduced solution obtained in step (5) to increase. 2+ In addition, since less S is generated in reaction (5), after returning to the second stage oxygen pressure leaching, the reaction progress of reaction (6) and the heat generated are limited, thus resulting in a significant increase in the amount of steam used in the second stage oxygen pressure leaching process.

[0162] Comparative Example 8

[0163] The method for removing iron from zinc by zinc oxygen pressure leaching in this comparative example comprises the following steps:

[0164] (1) Zinc concentrate (particle size -44 μm, solid content greater than 97%, Zn 49%, Fe 8%, S29%, Cu 0.5%, specific surface area 3.75 m 2 / cm 3 )400g, neutralized slag (16.33g, Zn 53%, Fe 9%, S1.5%) were mixed, lignin (1.6g) was added, iron removal solution (1L, H2SO4 40g / L, Zn 97.6g / L, Fe 1.6g / L, Cu 0.28g / L), second immersion solution (1.4L, H2SO4 39g / L, Zn109.15g / L, Fe 10.92g / L, Cu 0.76 g / L), oxygen gas with an oxygen concentration of 99 vol% was introduced for a stage of oxygen pressure leaching, the reaction temperature was 150°C, the total pressure was 1.25 MPa, the reaction time was 1.5 h, and after cooling and pressure reduction and liquid-solid separation, a leaching residue (275 g, Zn10.85%, Fe22.65%, Cu0.3%, S40.86%) and a leaching solution (2.4 L, H2SO4 12.6 g / L, Zn 173.92 g / L, Fe (total iron) 1.7 g / L, Fe 2+ 0.5g / L, Cu 1.16g / L), the first immersion liquid is sent to the subsequent pre-neutralization, purification, electrolysis and casting processes.

[0165] (2) All the above-mentioned leaching slag and reduction slag (184 g, Zn 43%, Fe 7.7%, Cu 0.52%, S 30%, specific surface area 4.42 m 2 / cm 3 ) was mixed, lignin (1.6 g) was added, 1.4 L of spent electrolyte (Zn 43 g / L, H2SO4 160 g / L) was added, at this time, the acid-zinc molar ratio was controlled to be 1.37, oxygen gas with an oxygen concentration of 99 vol% was introduced for two-stage oxygen pressure leaching, the reaction temperature was controlled to be 150 ° C and the total pressure was 1.25 MPa according to the steam introduction amount, and the reaction time was 2 h. After cooling, pressure reduction and liquid-solid separation, the second leaching residue (330.48 g, Zn 8.8%, Fe 21.01%, Cu 0.1%, S 51.63%) and the second leaching solution (1.4 L, H2SO4 39 g / L, Zn 109.15 g / L, Fe 10.92 g / L, Fe 2+ 0.9g / L, Cu 0.76g / L), wherein the steam consumption in the second stage oxygen pressure leaching process is 0.03t steam / t zinc concentrate.

[0166] (3) returning all the secondary leaching liquid to the first oxygen pressure leaching stage;

[0167] 1L of spent electrolyte (Zn 43g / L, H2SO4 160g / L) was added to the above-mentioned second-stage slag (330.48g, Zn 8.8%, Fe 21.01%, Cu 0.1%, S 51.63%), and reacted at 92℃ for 2h, followed by solid-liquid separation to obtain hot acid solution (1L, H2SO4 40g / L, Zn 66.27g / L, Fe 56.94g / L, Fe 2+ 0.8g / L, Cu 0.20g / L) and hot acid slag (264.38g, Zn4.2%, Fe5%, Cu 0.08%, S 90.2%), and the hot acid slag is sent to the flotation process for further treatment.

[0168] (4) Add the hot acid solution (1L, H2SO4 40g / L, Zn 66.27g / L, Fe 56.94g / L, Fe 2+ 0.8 g / L, Cu0.20 g / L) was slowly added with 32.65 g of zinc roasted sand (Zn 57%, Fe 8.5%, S1%, Cu 0.55%) at 85°C, and the reaction was continued for 75 min. The endpoint pH was controlled to be 2.5. The neutralized slag (16.33 g, Zn 53%, Fe 9%, Cu 0.57%, S1.5%) and the neutralized solution (1 L, Zn 83.95 g / L, Fe 58.07 g / L, Fe 2+ 0.7g / L, Cu 0.28g / L).

[0169] (5) returning the neutralized slag to a first stage for oxygen pressure leaching;

[0170] The above neutralization solution (1L, Zn 83.95g / L, Fe58.07g / L, Fe 2+ 0.7g / L, Cu 0.28g / L) for reduction treatment, wherein 200g of zinc concentrate was added, the reaction temperature was 85°C, the reaction time was 300min, and after the reaction was completed, liquid-solid separation was performed to obtain reduction slag (184g, Zn43%, Fe 7.7%, Cu 0.52%, S30%) and reduction solution (1L, Zn89.83g / L, Fe59.24g / L, Fe 2+ 58.94 g / L, Cu 0.28 g / L).

[0171] (6) returning the reduced slag (184 g, Zn 43%, Fe 7.7%, Cu 0.52%, S 30%) to the second-stage oxygen pressure leaching process;

[0172] The above reduced solution (1L, Zn89.83g / L, Fe59.24g / L, Fe 2+58.94g / L, Cu 0.28g / L) was added into the autoclave, and oxygen with a purity of 99vol% was introduced at 190℃, and the total pressure was controlled at 1.8MPa. After the reaction for 300min, the temperature and pressure were lowered to obtain iron removal liquid (1L, Zn 89.78g / L, Fe1.6g / L, Cu 0.28g / L, H2SO4 40g / L) and iron slag (101.34g, Zn0.2%, Fe 56.75%, S1.2%), and the iron slag was sent to subsequent resource recovery. Among them, the steam consumption of the iron removal process is 1.05t steam / t zinc concentrate.

[0173] (7) Returning the iron removal solution to the first oxygen pressure leaching process.

[0174] It is calculated that in the above whole process, the recovery rate of Zn is 96.1%; the recovery rate of iron is 66%.

[0175] The results show that increasing the amount of zinc concentrate added in step (5) to 200 g, and adding too much zinc concentrate, can satisfy the requirement of Fe 2+ However, due to the excessive amount of unreacted zinc concentrate remaining, the overall specific surface area of ​​the reduced slag is reduced, which affects the heat supply per unit time of the reaction (6). The steam consumption in the second-stage oxygen pressure leaching process has increased to a certain extent. In addition, due to the excessive amount of unreacted zinc sulfide, the zinc content in the second leaching slag increases, resulting in a decrease in the total recovery rate of zinc. The increase in the total amount of zinc concentrate leads to a decrease in the iron resource utilization rate.

[0176] 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 removing iron by zinc oxygen pressure leaching, characterized in that: The steps include: S1. After the zinc concentrate to be treated is subjected to a stage of oxygen pressure leaching, solid-liquid separation is performed to obtain a leaching residue and a leaching liquid; Wherein, the zinc concentrate is zinc sulfide concentrate; S2, subjecting the first leaching residue to two-stage oxygen pressure leaching, and then separating the solid and liquid to obtain a second leaching residue and a second leaching liquid; S3, returning the secondary leaching solution to the first oxygen pressure leaching process in S1; After hot acid leaching of the secondary leaching residue, solid-liquid separation is performed to obtain hot acid residue and Fe-rich 3+ of hot acid; S4, after neutralizing the hot acid liquid, separate the solid and liquid to obtain a neutralized slag and a neutralized liquid with a pH value of 2 to 4; S5, 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; S6, 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 second-stage oxygen pressure leaching process in S2; S7, returning the iron removal solution to the oxygen pressure leaching process in S1.

2. The method according to claim 1, characterized in that During the first stage of oxygen pressure leaching, oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced, the initial acid concentration of the reaction system is controlled to be 30-60 g / L, the reaction temperature is 140-160° C., the total pressure is 1.0-1.3 MPa, and the reaction time is 1-2 h; Preferably, during the first stage of oxygen pressure leaching, a surfactant is added to the reaction system, wherein the amount of the surfactant added is 3-5‰ of the zinc concentrate; preferably, the surfactant is lignin.

3. The method according to claim 1 or 2, characterized in that: When performing the second stage oxygen pressure leaching, an oxygen-rich gas with an oxygen concentration of ≥50 vol% is introduced, and a first solution containing sulfuric acid is added, the initial molar ratio of acid to zinc in the reaction system is controlled to be 2-3:1, the reaction temperature is 140-160°C, the total pressure is 1.0-1.3 MPa, and the reaction time is 1.5-3h; preferably, when performing the second stage oxygen pressure leaching, a surfactant is added to the reaction system, wherein the amount of the surfactant added is 3-5‰ of the zinc concentrate corresponding to the first leaching residue; preferably, the surfactant is lignin; Preferably, the final acid concentration of the second immersion solution is 35 to 65 g / L; Preferably, the concentration of sulfuric acid in the first solution is 140-180 g / L, and more preferably, the first solution is a waste electrolyte.

4. The method according to claim 3, characterized in that In S3, when hot acid leaching is performed, the leaching temperature is controlled to be 80 to 99° C., preferably 90 to 95° C.; Preferably, the secondary 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 secondary leaching residue to the second solution is 1 g: 2-5 mL; More preferably, the second solution is a waste electrolyte; Preferably, the amount of the second solution added in S3 is 35-45 vol% of the total amount of the first solution in S2 and the second solution in S3.

5. The method according to claim 1, characterized in that In S4, 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 60-90 minutes.

6. The method according to claim 1, characterized in that In S5, when the reduction treatment is performed, the amount of the reducing agent added is such that Fe 3+ Completely converted to Fe 2+ 1.2-1.8 times, preferably 1.35-1.65 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.

7. The method according to claim 1, characterized in that In S6, 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.

8. The method according to any one of claims 1 to 7, 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 content of Zn is 48-51%, the content of Fe is 6-9%, the content of S is 28-32%, and the content of Cu is 0.4-0.7%.

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