A combined device and method for floating and smelting lead and zinc in a refractory complex ore

Abstract

The application discloses a combined device and method for flotation-metallurgy of lead and zinc in refractory complex ores, comprising a flotation system and an oxygen pressure acid leaching system; the flotation system comprises floaters connected in sequence; the oxygen pressure acid leaching system comprises a first-stage oxygen pressure acid leaching system and a second-stage oxygen pressure acid leaching system, a desulfurization reactor and a lead selection reactor; the desulfurization reactor comprises a rotatable open tank cover and a solid-liquid separator; the lead selection reactor comprises a temperature control system and a solid-liquid separation pipe. The steps of using the device are as follows: the raw ore is mixed and floated to separate lead and zinc in the flotation system, and the mixed lead-zinc concentrate is sent to the oxygen pressure acid leaching system to extract lead and zinc respectively. The application has the characteristics of good metal recovery effect, clear and smooth process flow, convenient operation and environmental friendliness.

Description

TECHNICAL FIELD

[0001] The application belongs to the technical field of mineral processing, and relates to a combined device and method for flotation-metallurgy of lead and zinc in a refractory complex ore. BACKGROUND

[0002] The current non-ferrous metal in China has the characteristics of more lean ore and less rich ore, more associated ore and less single ore. The processing method for lead-zinc ore still adopts the traditional preferential flotation method, and three products of lead concentrate, zinc concentrate and tailings are obtained respectively, and then metal refining is carried out, which leads to complex production process, low recovery rate of lead and zinc, many types of added reagents, complex tail water composition and difficult treatment. At present, the processing of refractory lead-zinc ore is mostly separated into ore dressing and smelting, and more consideration is given to obtaining high-grade concentrate by using ore dressing technology, and less consideration is given to reasonably connecting and optimizing the process of ore dressing and smelting, so that the ore dressing process is complex, and the product after ore dressing needs to be transferred to the next reaction area, which consumes manpower and material resources, is low in efficiency, and has low metal recovery rate. SUMMARY

[0003] In order to overcome the shortcomings of the prior art, the application provides a combined device and method for flotation-metallurgy of lead and zinc in a refractory complex ore, which has good metal recovery effect, clear and smooth process flow, is convenient to operate and environment-friendly.

[0004] In order to achieve the above purpose, the technical scheme of the application is as follows:

[0005] A combined device for flotation-metallurgy of lead and zinc in a refractory complex ore, comprising a flotation system and an oxygen pressure acid leaching system; the oxygen pressure acid leaching system comprises a first-stage oxygen pressure acid leaching system, a second-stage oxygen pressure acid leaching system, a desulfurization reactor, a lead selection reaction tank, a purification tank and a zinc electrolysis tank;

[0006] The first-stage oxygen pressure acid leaching system comprises a ball mill, a first slurry adjusting tank, a first-stage oxygen pressure kettle, a first-stage flash tank, a first-stage intermediate tank and a first-stage filter press connected in sequence;

[0007] The second-stage oxygen pressure acid leaching system comprises a second slurry adjusting tank, a second-stage oxygen pressure kettle, a second-stage flash tank, a second-stage intermediate tank and a second-stage filter press connected in sequence;

[0008] The liquid outlet of the first-stage filter press is connected with the purification tank, the slag outlet of the first-stage filter press is connected with the second slurry adjusting tank, the liquid outlet of the zinc electrolysis tank is connected with the second slurry adjusting tank, the liquid outlet of the second-stage filter press is connected with the first slurry adjusting tank, the slag outlet of the second-stage filter press is connected with the desulfurization reactor, and the slag outlet of the desulfurization reactor is connected with the lead selection reaction tank.

[0009] As a further technical improvement, the flotation system comprises a plurality of flotation devices connected in sequence, the flotation device comprising a base, an overflow tank, a flotation motor and an impeller, a flotation zone is formed between the base and the overflow tank, the lower end of the flotation zone is communicated with an air charging pipe, and the other end of the air charging pipe is communicated with an air charging pump; the flotation motor is arranged at the center of the base and connected with the impeller through a rotating shaft extending out of the base, and the overflow tank is arranged around the top of the flotation zone;

[0010] A concentrate conveying pipe is arranged below the overflow tank and communicated with the flotation zone of the next stage flotation device;

[0011] A middling inlet and a tailing outlet are arranged at the lower part of the two sides of the flotation zone respectively, the tailing outlet of the next stage flotation device is communicated with the middling inlet of the previous stage flotation device through a tailing conveying pipe, and a peristaltic pump is installed on the tailing conveying pipe.

[0012] As a further technical improvement, the flotation system comprises a roughing flotation device, the roughing flotation device is connected with a cleaning one flotation device and a cleaning two flotation device respectively, the cleaning one flotation device is connected with a cleaning three flotation device in sequence, the cleaning two flotation device is connected with a cleaning four flotation device in sequence, and the cleaning three flotation device is communicated with a ball mill through a concentrate conveying pipe.

[0013] As a further technical improvement, the desulfurization reactor comprises a reaction tank, a tank cover, a solid-liquid separator, a rotating cover opening motor and a double motor controller;

[0014] The top of the reaction tank is connected with the tank cover, and the tank cover is rotatably connected with the reaction tank in parallel, and the rotating shaft is fixed with the upper port edge of the reaction tank;

[0015] The lower end of the reaction tank is in the shape of an inverted cone, and the bottom is connected with the solid-liquid separator, the solid-liquid separator is rotatably connected with the reaction tank in parallel, and the rotating shaft is fixed with the lower port edge of the reaction tank; the rotating shafts of the tank cover and the solid-liquid separator are respectively connected with a rotating cover opening motor, the rotating cover opening motor is connected with the double motor controller, and the double motor controller is arranged at the lower part of the tank body;

[0016] The upper end of the solid-liquid separator is a desulfurization reactor filter screen, and the lower end is a desulfurization reactor slag outlet, and a desulfurization reactor liquid outlet pipe is arranged at the side of the lower part of the solid-liquid separator.

[0017] As a further technical improvement, the lead selection reaction tank is provided with a temperature control mechanism and a solid-liquid separation pipe, the temperature control mechanism comprises a fixed seat, an electric heating wire, a heat dissipation fan, a heating motor, a heat dissipation motor and a temperature control motor controller, the electric heating wire is arranged on both sides of the inside of the fixed seat, the lead selection reaction tank is arranged in the fixed seat, the heat dissipation fan is arranged at the lower end of the electric heating wire, the electric heating wire is connected with the heating motor, the heat dissipation fan is connected with the heat dissipation motor, the heating motor and the heat dissipation motor are arranged on the fixed seat and connected with the temperature control motor controller, and the temperature control motor controller is arranged on the lower part of the outside of the fixed seat.

[0018] As a further technical improvement, the first and second pulp conditioning tanks are provided with temperature control devices to adjust the reaction temperature in the tanks.

[0019] As a further technical improvement, the use method of the combined device for flotation-metallurgy of lead and zinc in a refractory complex ore comprises the following steps:

[0020] A. Grinding: The raw ore is ground to a fineness of 70-80% passing 200 mesh and is conveyed to the flotation system.

[0021] B. Flotation of lead and zinc:

[0022] The following reagents are added respectively and the flotation is carried out according to the steps:

[0023] ① Roughing: 15 kg / t of lime, 200 g / t of ethion, 200 g / t of isobutyl sodium (potassium) xanthate, 300 g / t of sodium hexametaphosphate and 70 g / t of pine oil are added; the flotation motor is started, the concentrate froth of the roughing overflow tank is transported to the first cleaning flotation tank, and the tailings are transported to the first scavenging flotation tank through the tailings conveying pipe;

[0024] The ethion and isobutyl sodium (potassium) xanthate are collectors, the ethion has strong collecting ability, fast flotation speed and low reagent consumption, and its collecting ability for pyrite is very weak; the isobutyl sodium (potassium) xanthate is mainly used for the flotation of various metal sulfide ores such as copper, lead and zinc, but it is particularly effective for the flotation of pyrite;

[0025] The lime is an inhibitor and a pH regulator, which functions to inhibit pyrite and adjust the pH of the ore slurry to be weakly alkaline;

[0026] The sodium hexametaphosphate is an inhibitor, which can combine with calcium and magnesium ions to form a complex and prevent the floating of gangue minerals;

[0027] The pine oil is a frother.

[0028] ② Scavenging:

[0029] Scavenging I process adds lime 3500g / t, ethiofencarb 200g / t, sodium (potassium) isobutyl xanthate 200g / t, sodium hexametaphosphate 300g / t, copper sulfate 300g / t, pine oil 50g / t; start the flotation motor, the concentrate foam of the overflow tank of the scavenging I flotation cell is transported back to the roughing flotation cell, and the tailings are transported to the scavenging II flotation cell through the tailings transport pipe;

[0030] The copper sulfate is an activator, which makes the collector more easily collect the activated sphalerite.

[0031] Scavenging II process adds lime 2000g / t, ethiofencarb 100g / t, sodium (potassium) isobutyl xanthate 100g / t, sodium hexametaphosphate 150g / t, copper sulfate 150g / t, pine oil 30g / t; start the flotation motor, the concentrate foam of the overflow tank of the scavenging II flotation cell is transported back to the scavenging I flotation cell, and the tailings are transported to the scavenging III flotation cell through the tailings transport pipe;

[0032] Scavenging III process adds ethiofencarb 50g / t, sodium (potassium) isobutyl xanthate 50g / t, sodium hexametaphosphate 50g / t, copper sulfate 50g / t, pine oil 10g / t; start the flotation motor, the concentrate foam of the overflow tank of the scavenging III flotation cell is transported back to the scavenging II flotation cell, and the tailings are collected and treated through the tailings transport pipe.

[0033] ③Cleaning:

[0034] Cleaning I process adds lime 1500g / t, ethiofencarb 200g / t, sodium (potassium) isobutyl xanthate 200g / t; start the flotation motor, the concentrate foam of the overflow tank of the cleaning I flotation cell is transported to the cleaning II flotation cell, and the tailings are transported back to the roughing flotation cell through the tailings transport pipe.

[0035] Cleaning II process adds lime 1000g / t, ethiofencarb 150g / t, sodium (potassium) isobutyl xanthate 150g / t; start the flotation motor, the concentrate foam of the overflow tank of the cleaning II flotation cell is transported to the cleaning III flotation cell, and the tailings are transported back to the cleaning I flotation cell through the tailings transport pipe.

[0036] Cleaning III process adds lime 500g / t, ethiofencarb 60g / t, sodium (potassium) isobutyl xanthate 60g / t; start the flotation motor, the mixed lead-zinc concentrate produced by the overflow tank of the cleaning III flotation cell is transported to the I-stage oxygen pressure acid leaching system through the concentrate transport pipe, and the tailings are transported back to the cleaning II flotation cell through the tailings transport pipe.

[0037] C.Pressurized oxidation acid leaching

[0038] ①Slurry preparation:

[0039] The mixed lead-zinc concentrate is mixed with a sulfuric acid solution with a mass concentration of 140 g / L at a solid-liquid mass ratio of 1:(3-5) and is fed into a ball mill to be ground to a particle size of -200-100 mesh, is fed into a first pulp conditioning tank, and sodium lignosulfonate is added to obtain a pulp conditioning solution, wherein the sodium lignosulfonate is added in an amount of 0.2-0.5 wt% of the mixed lead-zinc concentrate.

[0040] ② Stage I pressure oxidation acid leaching:

[0041] After being preheated to 80-120°C by the first pulp conditioning tank, the stage I pressure oxidation autoclave is added, the temperature is raised to 150-180°C, oxygen is introduced to pressurize to 0.6-2.0 MPa, stage I pressure oxidation acid leaching is performed, the reaction time is 60-180 min, then after being flashed and depressurized by a stage I flash tank, the stage I pressure oxidation autoclave is fed into a stage I intermediate tank, and then after being solid-liquid separated by a stage I filter press, stage I leaching solution and stage I leaching residue are obtained.

[0042] ③ Stage II pressure oxidation acid leaching

[0043] The stage I leaching residue and waste electrolyte after electro-deposition of the stage I leaching solution are added into a second pulp conditioning tank for pulp conditioning at a liquid-solid mass ratio of (4-10):1, preheated to 80-120°C, and then added into a stage II pressure oxidation autoclave, oxygen is introduced to pressurize to 0.4-1.2 MPa, stage II pressure oxidation acid leaching is performed, the reaction time is 100-360 min, then after being flashed and depressurized by a stage II flash tank, the stage II pressure oxidation autoclave is fed into a stage II intermediate tank, and then after being solid-liquid separated by a stage II filter press, stage II leaching solution and stage II leaching residue are obtained, and the stage II leaching solution is returned to the first pulp conditioning tank to continue to participate in stage I pressure oxidation acid leaching;

[0044] The purpose of this step is to maximize the enrichment of zinc, copper and other elements in the stage I leaching residue into the stage II leaching solution, and to return to the stage I pressure oxidation acid leaching to further realize the enrichment and separation of zinc; at the same time, lead and sulfur are enriched into the stage II leaching residue.

[0045] ④ Stage II leaching residue treatment:

[0046] Desulfurization: the stage II leaching residue is fed into a desulfurization reactor, 4 mol / L of ammonium sulfide solution is added at a liquid-solid volume mass ratio of 11 mL / g, and the reaction time is maintained at room temperature for 40-50 min, and the sulfur leaching rate can reach 98%; the filtrate is separated, and sulfur can be obtained by thermal decomposition of the filtrate.

[0047] Lead selection: the filter residue in the desulfurization process is fed into a lead selection reaction tank, 7 mol / L of nitric acid solution is added at a liquid-solid ratio of 4:1, the leaching time is 50 min, and the leaching temperature is 80°C; after cooling, filtration and washing, the filtrate is added with sulfuric acid to make lead sulfate precipitate, filtration is performed, lead is separated, lead sulfate is converted into lead acetate, and then lead is obtained by replacement.

[0048] ④ Stage I leaching solution treatment:

[0049] The leaching solution from Section I is first discharged into the purification tank and then into the zinc electrolytic cell. After separation, purification and electrolysis, waste electrolyte and electrolytic zinc are obtained respectively. The waste electrolyte returns to the intermediate tank of Section I and continues to pass through the filter press of Section I for solid-liquid separation. The electrolytic zinc is then melted and cast to obtain zinc ingots.

[0050] Compared with the prior art, the present invention has the following advantages and beneficial effects:

[0051] 1. The present invention uses a combined beneficiation and smelting process, which can effectively recycle and utilize difficult-to-select lead-zinc ores and effectively separate lead and zinc compared to single flotation or hydrometallurgy.

[0052] 2. The present invention adopts the fully wet zinc smelting method, which has low unit energy consumption and is environmentally friendly.

[0053] 3. The device of the present invention has a high degree of integration, a clear and simple process, and high working efficiency.

[0054] 4. The desulfurization reactor in the present invention can electrically control the lid to be closed or opened, and the rotation of the solid-liquid separator is controlled by a motor to effectively achieve solid-liquid separation.

[0055] 5. The lead selection reaction tank of the present invention can electrically control the temperature in the tank, thereby ensuring the temperature conditions of the reaction and being easy and convenient to operate. BRIEF DESCRIPTION OF THE DRAWINGS

[0056] Figure 1 The figure is a schematic structural diagram of a combined device for flotation and smelting lead and zinc in difficult-to-select complex ores according to the present invention.

[0057] Figure 2 It is a structural schematic diagram of the flotation device of the present invention.

[0058] Figure 3 It is a schematic diagram of the upper and lower connections of the flotation device of the present invention.

[0059] Figure 4 It is a structural schematic diagram of the desulfurization reactor of the present invention.

[0060] Figure 5 This is a schematic structural diagram of the lead selection reaction tank of the present invention.

[0061] Figure ID:

[0062] 1 - 1st stage oxygen pressure acid leaching system, 11 - ball mill, 12 - 1st stage pulp conditioning tank, 13 - 1st stage oxygen pressure tank, 14 - 1st stage flash tank, 15 - 1st stage intermediate tank, 16 - 1st stage filter press, 161 - 1st stage filter press liquid outlet, 162 - 1st stage filter press slag outlet, 2 - 2nd stage oxygen pressure acid leaching system, 21 - 2nd stage pulp conditioning tank, 22 - 2nd stage oxygen pressure tank, 23 - 2nd stage flash tank, 24 - 2nd stage intermediate tank, 25 - 2nd stage filter press, 251 - 2nd stage filter press liquid outlet, 252 - 2nd stage filter press slag outlet, 3 - desulfurization reactor, 31 - reaction tank, 32 - tank cover, 33 - solid-liquid separator, 331 - desulfurization reactor filter screen, 332 - desulfurization reactor slag outlet, 333 - desulfurization reactor liquid outlet pipe, 34 - rotary cover opening motor, 35 - double motor controller, 36 - rotating shaft, 4 - lead selection reaction tank, 41 - temperature control mechanism, 411 - fixed seat, 412 - electric heating wire, 413 - heat dissipation fan, 414 - heating motor, 415 - heat dissipation motor, 416 - temperature control motor controller, 42 - solid-liquid separation pipe, 5 - purification tank, 6 - zinc electrolysis tank, 61 - zinc electrolysis tank liquid outlet, 7 - flotation system, 71 - roughing flotation cell, 72 - first cleaning flotation cell, 73 - second cleaning flotation cell, 74 - third cleaning flotation cell, 75 - first scavenging flotation cell, 76 - second scavenging flotation cell, 77 - third scavenging flotation cell, 8 - flotation cell, 81 - base, 82 - overflow tank, 83 - flotation motor, 84 - impeller, 85 - flotation zone, 851 - air charging pipe, 852 - air charging pump, 86 - concentrate delivery pipe, 87 - middlings inlet, 88 - tailings outlet, 89 - tailings delivery pipe, 891 - peristaltic pump. DETAILED DESCRIPTION

[0063] The application will be further described below with reference to the drawings. Example 1

[0064] A combined lead-zinc flotation-metallurgy device for a refractory complex ore includes a flotation system 7 and an oxygen pressure acid leaching system; the oxygen pressure acid leaching system includes a 1st stage oxygen pressure acid leaching system 1, a 2nd stage oxygen pressure acid leaching system 2, a desulfurization reactor 3, a lead selection reaction tank 4, a purification tank 5, and a zinc electrolysis tank 6.

[0065] The 1st stage oxygen pressure acid leaching system 1 includes a ball mill 11, a 1st stage pulp conditioning tank 12, a 1st stage oxygen pressure tank 13, a 1st stage flash tank 14, a 1st stage intermediate tank 15, and a 1st stage filter press 16 connected in sequence.

[0066] The 2nd stage oxygen pressure acid leaching system 2 includes a 2nd stage pulp conditioning tank 21, a 2nd stage oxygen pressure tank 22, a 2nd stage flash tank 23, a 2nd stage intermediate tank 24, and a 2nd stage filter press 25 connected in sequence.

[0067] The liquid outlet 161 of the first pressure filter is connected with the purification tank 5, the slag outlet 162 of the first pressure filter is connected with the second pulp adjusting tank 21, the liquid outlet 61 of the zinc electrolysis tank is connected with the second pulp adjusting tank 21, the liquid outlet 251 of the second pressure filter is connected with the first pulp adjusting tank 12, the slag outlet 252 of the second pressure filter is connected with the desulfurization reactor 3, and the slag outlet 332 of the desulfurization reactor is communicated with the lead selection reaction tank 4;

[0068] The method for grinding and beneficiation of the refractory lead-zinc sulfide ore using the above device specifically comprises the following steps:

[0069] A. Grinding: the raw ore is ground to a fineness of 70% passing 200 mesh, and is transported to the flotation system 7.

[0070] B. Flotation of lead and zinc:

[0071] The following reagents are added respectively, and the flotation is carried out according to the steps:

[0072] ① Roughing: 15 kg / t of lime, 200 g / t of ethylthiuram, 200 g / t of sodium (potassium) isobutyl xanthate, 300 g / t of sodium hexametaphosphate and 70 g / t of pine oil are added; the flotation motor 83 is started, and the concentrate froth of the overflow tank of the roughing flotation device 71 is transported to the cleaning flotation device 72, and the tailings are transported to the scavenging flotation device 75 through the tailings conveying pipe 89;

[0073] Ethylthiuram and sodium (potassium) isobutyl xanthate are collectors, and ethylthiuram has strong collecting ability, fast flotation speed and low reagent consumption, and its collecting ability for pyrite is very weak; sodium (potassium) isobutyl xanthate is mainly used for the flotation of various metal sulfide ores such as copper, lead and zinc, but it is particularly effective for the flotation of pyrite;

[0074] Lime is an inhibitor and a pH regulator, and its function is to inhibit pyrite and adjust the pH of the ore slurry to be weakly alkaline;

[0075] Sodium hexametaphosphate is an inhibitor, which can combine with calcium and magnesium ions to form a complex, thereby preventing the floating of gangue minerals;

[0076] Pine oil is a frother.

[0077] ② Scavenging:

[0078] In the process of scavenging I, 3500 g / t of lime, 200 g / t of ethylthiuram, 200 g / t of sodium (potassium) isobutyl xanthate, 300 g / t of sodium hexametaphosphate, 300 g / t of copper sulfate and 50 g / t of pine oil are added; the flotation motor 83 is started, and the concentrate froth of the overflow tank of the scavenging flotation device 75 is transported back to the roughing flotation device 71, and the tailings are transported to the scavenging flotation device 76 through the tailings conveying pipe 89;

[0079] Copper sulfate is an activator, which makes the collector more easily collect the activated sphalerite.

[0080] The sweep III process adds ethylthiourea 50 g / t, sodium (potassium) isobutyl xanthate 50 g / t, sodium hexametaphosphate 50 g / t, copper sulfate 50 g / t, and pine oil 10 g / t; the flotation motor 83 is started, and the concentrate froth from the overflow tank of the sweep three flotation cell 77 is transported back to the sweep two flotation cell 76, and the tailings are collected and treated through the tailings transport pipe 89.

[0081] The sweep III process adds ethylthiourea 50 g / t, sodium (potassium) isobutyl xanthate 50 g / t, sodium hexametaphosphate 50 g / t, copper sulfate 50 g / t, and pine oil 10 g / t; the flotation motor 83 is started, and the concentrate froth from the overflow tank of the sweep three flotation cell 77 is transported back to the sweep two flotation cell 76, and the tailings are collected and treated through the tailings transport pipe 89.

[0082] ③ Concentration:

[0083] The concentration I process adds lime 1500 g / t, ethylthiourea 200 g / t, and sodium (potassium) isobutyl xanthate 200 g / t; the flotation motor 83 is started, and the concentrate froth from the overflow tank of the concentration one flotation cell 72 is transported to the concentration two flotation cell 73, and the tailings are transported back to the rough flotation cell 71 through the tailings transport pipe 89.

[0084] The concentration II process adds lime 1000 g / t, ethylthiourea 150 g / t, and sodium (potassium) isobutyl xanthate 150 g / t; the flotation motor 83 is started, and the concentrate froth from the overflow tank of the concentration two flotation cell 73 is transported to the concentration three flotation cell 74, and the tailings are transported back to the concentration one flotation cell 72 through the tailings transport pipe 89.

[0085] The concentration III process adds lime 500 g / t, ethylthiourea 60 g / t, and sodium (potassium) isobutyl xanthate 60 g / t; the flotation motor 83 is started, and the mixed lead-zinc concentrate produced by the overflow tank of the concentration three flotation cell 74 is transported to the I-stage oxygen pressure acid leaching system 1 through the concentrate transport pipe 86, and the tailings are transported back to the concentration two flotation cell 73 through the tailings transport pipe 89.

[0086] C. Pressure oxidation acid leaching

[0087] ① Slurry preparation:

[0088] The mixed lead-zinc concentrate is mixed with a sulfuric acid solution with a mass concentration of 140 g / L at a solid-liquid mass ratio of 1:3 and transported into a ball mill 11 to a particle size of -200 mesh, input into a first slurry preparation tank 12, and sodium lignosulfonate is added to obtain a slurry, and the addition amount of sodium lignosulfonate is 0.2wt% of the mixed lead-zinc concentrate.

[0089] ② I-stage pressure oxidation acid leaching:

[0090] After preheating to 80℃, the first slurry tank 12 is added to the first oxygen pressure tank 13, heated to 150℃, and oxygen is introduced to pressurize to 0.6MPa, and the first stage of pressure oxidation acid leaching is carried out, the reaction time is 60min, then after flash decompression through the first stage flash tank 14 into the first stage intermediate tank 15, and then after solid-liquid separation through the first stage filter press 16, the first stage leaching solution and the first stage leaching residue are obtained.

[0091] ③The second stage of pressure oxidation acid leaching

[0092] The first stage leaching residue and the waste electrolyte after the first stage leaching solution is electrolyzed are added to the second slurry tank 21 according to the liquid-solid mass ratio of 4:1 for slurry preparation, preheated to 80℃, then added to the second oxygen pressure tank 22, oxygen is introduced to pressurize to 0.4MPa, and the second stage of pressure oxidation acid leaching is carried out, the reaction time is 100min, then after flash decompression through the second stage flash tank 23 into the second stage intermediate tank 24, and then after solid-liquid separation through the second stage filter press 25, the second stage leaching solution and the second stage leaching residue are obtained, and the second stage leaching solution returns to the first slurry tank 12 to continue to participate in the first stage of pressure oxidation acid leaching;

[0093] The purpose of this step is to maximize the enrichment of zinc, copper and other elements in the first stage leaching residue into the second stage leaching solution, and to return to the first stage of pressure oxidation acid leaching to further realize the enrichment and separation of zinc; at the same time, lead and sulfur are enriched into the second stage leaching residue.

[0094] ④Treatment of the second stage leaching residue:

[0095] Desulfurization: The second stage leaching residue is input into the desulfurization reactor 3, and 4mol / L of ammonium sulfide solution is added according to the liquid-solid volume mass ratio of 11mL / g, and the reaction time is maintained at room temperature for 40min, and the sulfur leaching rate can reach 98%; the filtrate is separated, and sulfur can be obtained by thermal decomposition of the filtrate.

[0096] Lead selection: The filter residue in the desulfurization process is input into the lead selection reaction tank 4, and 7mol / L of nitric acid solution is added according to the liquid-solid ratio of 4:1, the leaching time is 50min, and the leaching temperature is 80℃, and after cooling, filtration and washing, the filtrate is added with sulfuric acid to make lead sulfate precipitate, and after filtration, lead is separated, and then lead sulfate is converted into lead acetate, and then replaced into lead.

[0097] ④Treatment of the first stage leaching solution:

[0098] The first stage leaching solution is first discharged into the purification tank 5, and then discharged into the zinc electrolysis tank 6, and after separation, purification and electrolysis, waste electrolyte and electrolytic zinc are obtained, respectively, wherein the waste electrolyte returns to the first stage intermediate tank 15, and continues to be separated by the first stage filter press 16, and the electrolytic zinc is melted to obtain zinc ingots. Example 2:

[0099] The difference from the embodiment 1 is that the flotation system comprises a plurality of flotation devices connected in sequence, the flotation system 7 comprises a plurality of flotation devices 8 connected in sequence, the flotation device 8 comprises a base 81, an overflow tank 82, a flotation motor 83 and an impeller 84, a flotation zone 85 is formed between the base 81 and the overflow tank 82, an air charging pipe 851 is communicated to the outside of the lower end of the flotation zone 85, and the other end of the air charging pipe 851 is communicated to an air charging pump 852; the flotation motor 83 is arranged at the center of the base 81 and connected to the impeller 84 through a rotating shaft 36 extending out of the base 81, and the overflow tank 82 is arranged around the top of the flotation zone 85;

[0100] A concentrate conveying pipe 86 is arranged below the overflow tank 82 and communicated to the flotation zone 85 of the next-stage flotation device 8;

[0101] A middling inlet 87 and a tailing outlet 88 are arranged at the lower part of the two sides of the flotation zone 85 respectively, the tailing outlet 88 of the next-stage flotation device 8 is communicated to the middling inlet 87 of the previous-stage flotation device 8 through a tailing conveying pipe 89, and a peristaltic pump 891 is installed on the tailing conveying pipe 89;

[0102] The method for grinding and beneficiation of a refractory lead-zinc sulfide ore using the above device comprises the following steps:

[0103] A. Grinding: the raw ore is ground to a fineness of -200 mesh 70-80% of the ore slurry, and is conveyed to the flotation system 7.

[0104] B. Flotation of lead and zinc:

[0105] The following reagents are added respectively, and the flotation is carried out according to the steps:

[0106] ① Roughing: 15 kg / t of lime, 200 g / t of ethylthiourea, 200 g / t of sodium (potassium) isobutyl xanthate, 300 g / t of sodium hexametaphosphate and 70 g / t of pine oil are added; the flotation motor 83 is started, the concentrate froth of the overflow tank of the roughing flotation device 71 is conveyed to the first-stage cleaning flotation device 72, and the tailings are conveyed to the first-stage scavenging flotation device 75 through the tailing conveying pipe 89;

[0107] The ethylthiourea and the sodium (potassium) isobutyl xanthate are collectors, the ethylthiourea has strong collecting ability, fast flotation speed and low reagent consumption, and has weak collecting ability for pyrite; the sodium (potassium) isobutyl xanthate is mainly used for the flotation of various metal sulfide ores such as copper, lead and zinc, but is particularly effective for the flotation of pyrite;

[0108] The lime is an inhibitor and a pH regulator, and functions to inhibit pyrite and adjust the pH of the ore slurry to be weak alkaline;

[0109] The sodium hexametaphosphate is an inhibitor, and can combine with calcium and magnesium ions to form a complex to prevent gangue minerals from floating;

[0110] The pine oil is a frother.

[0111] 2. Scavenging:

[0112] Scavenging I process adds lime 3500g / t, ethylthiourea 200g / t, sodium (potassium) isobutyl xanthate 200g / t, sodium hexametaphosphate 300g / t, copper sulfate 300g / t, pine oil 50g / t; start the flotation motor 83, the concentrate foam of the overflow tank of the scavenging I flotation cell 75 is transported back to the roughing flotation cell 71, and the tailings are transported to the scavenging II flotation cell 76 through the tailings transport pipe 89;

[0113] Copper sulfate is an activator, which makes the collector more easily collect the activated sphalerite.

[0114] Scavenging II process adds lime 2000g / t, ethylthiourea 100g / t, sodium (potassium) isobutyl xanthate 100g / t, sodium hexametaphosphate 150g / t, copper sulfate 150g / t, pine oil 30g / t; start the flotation motor 83, the concentrate foam of the overflow tank of the scavenging II flotation cell 76 is transported back to the scavenging I flotation cell 75, and the tailings are transported to the scavenging III flotation cell 77 through the tailings transport pipe 89;

[0115] Scavenging III process adds ethylthiourea 50g / t, sodium (potassium) isobutyl xanthate 50g / t, sodium hexametaphosphate 50g / t, copper sulfate 50g / t, pine oil 10g / t; start the flotation motor 83, the concentrate foam of the overflow tank of the scavenging III flotation cell 77 is transported back to the scavenging II flotation cell 76, and the tailings are collected and treated through the tailings transport pipe 89.

[0116] 3. Concentration:

[0117] Concentration I process adds lime 1500g / t, ethylthiourea 200g / t, sodium (potassium) isobutyl xanthate 200g / t; start the flotation motor 83, the concentrate foam of the overflow tank of the concentration I flotation cell 72 is transported to the concentration II flotation cell 73, and the tailings are transported back to the roughing flotation cell 71 through the tailings transport pipe 89.

[0118] Concentration II process adds lime 1000g / t, ethylthiourea 150g / t, sodium (potassium) isobutyl xanthate 150g / t; start the flotation motor 83, the concentrate foam of the overflow tank of the concentration II flotation cell 73 is transported to the concentration III flotation cell 74, and the tailings are transported back to the concentration I flotation cell 72 through the tailings transport pipe 89.

[0119] Concentration III process adds lime 500g / t, ethylthiourea 60g / t, sodium (potassium) isobutyl xanthate 60g / t; start the flotation motor 83, the mixed lead-zinc concentrate produced by the overflow tank of the concentration III flotation cell 74 is transported to the I-stage oxygen pressure acid leaching system 1 through the concentrate transport pipe 86, and the tailings are transported back to the concentration II flotation cell 73 through the tailings transport pipe 89.

[0120] C. Pressure oxidation acid leaching

[0121] ①Slurry preparation:

[0122] The mixed lead-zinc concentrate is mixed with a sulfuric acid solution with a mass concentration of 140 g / L at a solid-liquid mass ratio of 1:4 and is conveyed into the ball mill 11 to be ground to a particle size of -50 mesh. The ground product is input into the first slurry preparation tank 12, and sodium lignosulfonate is added to obtain a slurry solution. The amount of sodium lignosulfonate added is 0.3 wt% of the mixed lead-zinc concentrate.

[0123] ②First-stage pressure oxidation acid leaching:

[0124] After being preheated to 100°C in the first slurry preparation tank 12, the slurry is added into the first-stage oxygen pressure kettle 13, heated to 170°C, and pressurized to 1.4 MPa by introducing oxygen. First-stage pressure oxidation acid leaching is performed for 120 min. Then, the slurry is flashed and depressurized in the first-stage flash tank 14, and is input into the first-stage intermediate tank 15. After being separated by the first-stage pressure filter 16, the first-stage leaching solution and the first-stage leaching residue are obtained.

[0125] ③Second-stage pressure oxidation acid leaching

[0126] The first-stage leaching residue and the waste electrolyte after electro-deposition of the first-stage leaching solution are added into the second slurry preparation tank 21 at a liquid-solid mass ratio of 7:1 for slurry preparation. After being preheated to 100°C, the slurry is added into the second-stage oxygen pressure kettle 22, pressurized to 0.8 MPa by introducing oxygen, and subjected to second-stage pressure oxidation acid leaching for 250 min. Then, the slurry is flashed and depressurized in the second-stage flash tank 23, and is input into the second-stage intermediate tank 24. After being separated by the second-stage pressure filter 25, the second-stage leaching solution and the second-stage leaching residue are obtained. The second-stage leaching solution is returned to the first slurry preparation tank 12 to participate in the first-stage pressure oxidation acid leaching.

[0127] The purpose of this step is to maximize the enrichment of zinc, copper, and other elements in the first-stage leaching residue into the second-stage leaching solution, so as to return to the first-stage pressure oxidation acid leaching for further enrichment and separation of zinc; at the same time, lead and sulfur are enriched into the second-stage leaching residue.

[0128] ④Treatment of the second-stage leaching residue:

[0129] Desulfurization: The second-stage leaching residue is input into the desulfurization reactor 3, and 4 mol / L of ammonium sulfide solution is added at a liquid-solid volume mass ratio of 11 mL / g. The reaction is maintained at room temperature for 45 min, and the sulfur leaching rate can reach 98%. The filtrate is separated, and sulfur can be obtained by thermal decomposition of the filtrate.

[0130] Lead selection: The filter residue in the desulfurization process is input into the lead selection reaction tank 4, and 7 mol / L of nitric acid solution is added at a liquid-solid ratio of 4:1. The leaching time is 50 min, and the leaching temperature is 80°C. After cooling, the solution is filtered and washed. Sulfuric acid is added to the filtrate to precipitate lead sulfate. After filtration, lead is separated. Lead sulfate is converted into lead acetate, and then lead is obtained by displacement.

[0131] (4) The first stage leaching solution treatment:

[0132] The first stage leaching solution is first discharged into the purification tank 5, and then discharged into the zinc electrolysis tank 6, and after separation and purification and electrodeposition, waste electrolyte and zinc are obtained, wherein the waste electrolyte is returned to the first stage intermediate tank 15, and continues to pass through the first stage filter press 16 for solid-liquid separation, and the zinc is obtained after smelting to obtain zinc ingots. Example 3:

[0133] The difference from example 2 is that the flotation system 7 includes a roughing flotation device 71, the roughing flotation device 71 is connected with a cleaning one flotation device 72 and a cleaning two flotation device 75 on both sides respectively, the cleaning one flotation device 72 is connected with a cleaning two flotation device 73 and a cleaning three flotation device 74 in turn, the cleaning one flotation device 75 is connected with a cleaning two flotation device 76 and a cleaning three flotation device 77 in turn; the concentrate conveying pipe 86 of the cleaning three flotation device 74 is communicated with the ball mill 11;

[0134] The method for grinding and beneficiation of a refractory lead-zinc sulfide ore using the above device specifically includes the following steps:

[0135] A. Grinding: The raw ore is ground to a fineness of 80% passing 200 mesh, and is conveyed to the flotation system 7.

[0136] B. Flotation of lead and zinc:

[0137] The following reagents are added respectively, and the flotation is carried out according to the steps:

[0138] ① Roughing: 15 kg / t of lime, 200 g / t of ethyldithizone, 200 g / t of isobutyl xanthate sodium (potassium), 300 g / t of sodium hexametaphosphate, and 70 g / t of pine oil are added; the flotation motor 83 is started, and the concentrate froth of the roughing flotation device 71 overflow tank is transported to the cleaning one flotation device 72, and the tailings are transported to the cleaning one flotation device 75 through the tailings conveying pipe 89;

[0139] Ethyldithizone and isobutyl xanthate sodium (potassium) are collectors, ethyldithizone has strong collecting ability, fast flotation speed, and small reagent dosage, and its collecting ability for pyrite is very weak; isobutyl xanthate sodium (potassium) is mainly used for the flotation of multiple metal sulfide ores such as copper, lead, and zinc, but it is particularly effective for the flotation of pyrite;

[0140] Lime is an inhibitor and a pH regulator, and its role is to inhibit pyrite and adjust the pH of the ore slurry to be weakly alkaline;

[0141] Sodium hexametaphosphate is an inhibitor, which can combine with calcium and magnesium ions to form a complex, preventing gangue minerals from floating;

[0142] Pine oil is a frother.

[0143] ② Cleaning:

[0144] Scavenger 1 process adds lime 3500g / t, ethylthiourea 200g / t, sodium (potassium) isobutyl xanthate 200g / t, sodium hexametaphosphate 300g / t, copper sulfate 300g / t, pine oil 50g / t; start the flotation motor 83, the concentrate froth of the overflow launder of the scavenger 1 flotation cell 75 is transported back to the rougher flotation cell 71, the tailings are transported to the scavenger 2 flotation cell 76 via the tailings transport pipe 89;

[0145] Copper sulfate is an activator, which makes the collector more easily collect the activated sphalerite.

[0146] Scavenger 2 process adds lime 2000g / t, ethylthiourea 100g / t, sodium (potassium) isobutyl xanthate 100g / t, sodium hexametaphosphate 150g / t, copper sulfate 150g / t, pine oil 30g / t; start the flotation motor 83, the concentrate froth of the overflow launder of the scavenger 2 flotation cell 76 is transported back to the scavenger 1 flotation cell 75, the tailings are transported to the scavenger 3 flotation cell 77 via the tailings transport pipe 89;

[0147] Scavenger 3 process adds ethylthiourea 50g / t, sodium (potassium) isobutyl xanthate 50g / t, sodium hexametaphosphate 50g / t, copper sulfate 50g / t, pine oil 10g / t; start the flotation motor 83, the concentrate froth of the overflow launder of the scavenger 3 flotation cell 77 is transported back to the scavenger 2 flotation cell 76, the tailings are collected and treated via the tailings transport pipe 89.

[0148] ③Cleaning:

[0149] Cleaning 1 process adds lime 1500g / t, ethylthiourea 200g / t, sodium (potassium) isobutyl xanthate 200g / t; start the flotation motor 83, the concentrate froth of the overflow launder of the cleaner 1 flotation cell 72 is transported to the cleaner 2 flotation cell 73, the tailings are transported back to the rougher flotation cell 71 via the tailings transport pipe 89.

[0150] Cleaning 2 process adds lime 1000g / t, ethylthiourea 150g / t, sodium (potassium) isobutyl xanthate 150g / t; start the flotation motor 83, the concentrate froth of the overflow launder of the cleaner 2 flotation cell 73 is transported to the cleaner 3 flotation cell 74, the tailings are transported back to the cleaner 1 flotation cell 72 via the tailings transport pipe 89.

[0151] Cleaning 3 process adds lime 500g / t, ethylthiourea 60g / t, sodium (potassium) isobutyl xanthate 60g / t; start the flotation motor 83, the mixed lead-zinc concentrate produced by the overflow launder of the cleaner 3 flotation cell 74 is transported to the 1st stage oxygen pressure acid leaching system 1 via the concentrate transport pipe 86, the tailings are transported back to the cleaner 2 flotation cell 73 via the tailings transport pipe 89.

[0152] C.Pressurized oxidation acid leaching

[0153] ①Slurry adjustment:

[0154] The mixed lead-zinc concentrate is mixed with a sulfuric acid solution with a mass concentration of 140 g / L at a solid-liquid mass ratio of 1:5 and is conveyed into the ball mill 11 to be ground to a particle size of 100 mesh, is input into the first slurry adjustment tank 12, and sodium lignosulfonate is added to obtain a slurry adjustment solution. The amount of sodium lignosulfonate added is 0.5 wt% of the mixed lead-zinc concentrate.

[0155] ②First-stage pressure oxidation acid leaching:

[0156] After being preheated to 120°C by the first slurry adjustment tank 12, the mixed lead-zinc concentrate is added into the first-stage oxygen pressure kettle 13, is heated to 180°C, and is pressurized to 2.0 MPa by oxygen. First-stage pressure oxidation acid leaching is performed for 180 min. After being flash decompressed by the first-stage flash tank 14, the mixed lead-zinc concentrate is input into the first-stage intermediate tank 15 and is separated into a first-stage leaching solution and a first-stage leaching residue by the first-stage pressure filter 16.

[0157] ③Second-stage pressure oxidation acid leaching

[0158] The first-stage leaching residue and the waste electrolyte after electro-deposition of the first-stage leaching solution are added into the second slurry adjustment tank 21 at a liquid-solid mass ratio of 10:1 for slurry adjustment. After being preheated to 120°C, the mixed lead-zinc concentrate is added into the second-stage oxygen pressure kettle 22, is pressurized to 1.2 MPa by oxygen, and second-stage pressure oxidation acid leaching is performed for 360 min. After being flash decompressed by the second-stage flash tank 23, the mixed lead-zinc concentrate is input into the second-stage intermediate tank 24 and is separated into a second-stage leaching solution and a second-stage leaching residue by the second-stage pressure filter 25. The second-stage leaching solution is returned to the first slurry adjustment tank 12 to participate in the first-stage pressure oxidation acid leaching;

[0159] The purpose of this step is to maximize the enrichment of zinc, copper, and other elements in the first-stage leaching residue into the second-stage leaching solution, so as to return to the first-stage pressure oxidation acid leaching to further realize the enrichment and separation of zinc; at the same time, lead and sulfur are enriched into the second-stage leaching residue.

[0160] ④Treatment of the second-stage leaching residue:

[0161] Desulfurization: The second-stage leaching residue is input into the desulfurization reactor 3, 4 mol / L of ammonium sulfide solution is added at a liquid-solid volume mass ratio of 11 mL / g, and the reaction is maintained at room temperature for 50 min. The sulfur leaching rate can reach 98%. The filtrate is separated, and sulfur can be obtained by thermal decomposition of the filtrate.

[0162] Lead selection: The filter residue in the desulfurization process is input into the lead selection reaction tank 4, 7 mol / L of nitric acid solution is added at a liquid-solid ratio of 4:1, the leaching time is 50 min, and the leaching temperature is 80°C. After cooling, filtration and washing are performed. Sulfuric acid is added to the filtrate to make lead sulfate precipitate. After filtration, lead is separated. Lead sulfate is converted into lead acetate, and then lead is obtained by replacement.

[0163] ④Ⅰ stage leaching solution treatment:

[0164] The leaching solution from stage I is first discharged into the purification tank 5 and then into the zinc electrolytic tank 6. After separation, purification and electrolysis, waste electrolyte and electrolytic zinc are obtained respectively. The waste electrolyte returns to the intermediate tank 15 of stage I and continues to pass through the filter press 16 of stage I for solid-liquid separation. The electrolytic zinc is then melted and cast to obtain zinc ingots. Example 4:

[0165] The difference from Example 3 is that the desulfurization reactor 3 includes a reaction tank 31, a tank cover 32, a solid-liquid separator 33, a rotary cover opening motor 34 and a dual-motor controller 35;

[0166] The top of the reaction tank 31 is connected to a tank cover 32, which is rotatably connected to the reaction tank 31 in parallel, and a rotating shaft 36 is fixed to the edge of the upper end of the reaction tank 31;

[0167] The lower end of the reaction tank 31 is inverted conical, with a solid-liquid separator 33 connected to the bottom. The solid-liquid separator 33 is rotatably connected to the reaction tank 31 in parallel, and a rotating shaft 36 is fixed to the edge of the lower end of the reaction tank 31. The tank cover 32 and the rotating shaft 36 of the solid-liquid separator 33 are each connected to a rotary cover opening motor 34, which is connected to a dual-motor controller 35. The dual-motor controller 35 is located at the bottom of the tank body.

[0168] The upper end of the solid-liquid separator 33 is a desulfurization reactor filter screen 331 , and the lower end is a desulfurization reactor slag outlet 332 . A desulfurization reactor liquid outlet pipe 333 is provided on the side of the lower part of the solid-liquid separator 33 . Example 5:

[0169] The difference from Example 4 is that the lead selection reaction tank 4 is provided with a temperature control mechanism 41 and a solid-liquid separation tube 42. The temperature control mechanism 41 includes a fixed seat 411, an electric heating wire 412, a cooling fan 413, a heating motor 414, a cooling motor 415 and a temperature control motor controller 416. Electric heating wires 412 are arranged on both sides of the fixed seat 411. The lead selection reaction tank 4 is placed in the fixed seat 411. The cooling fan 413 is placed at the lower end of the electric heating wire 412. The electric heating wire 412 is connected to the heating motor 414. The cooling fan 413 is connected to the cooling motor 415. The heating motor 414 and the cooling motor 415 are arranged on the fixed seat 411 and connected to the temperature control motor controller 416. The temperature control motor controller 416 is arranged at the lower outer part of the fixed seat 411; the solid-liquid separation tube 42 is arranged at the front side of the lower end of the lead selection reaction tank 4. Example 6:

[0170] The difference from Example 5 is that the first slurry mixing tank 12 and the second slurry mixing tank 21 are provided with temperature control devices to adjust the reaction temperature in the tank.

[0171] Obviously, the above embodiments are only examples for clearly illustrating the present application, and are not intended to limit the implementation of the present application. Based on the above description, other different forms of changes or variations can be made by those skilled in the art; here, all the implementations do not need to be exhausted, and the obvious changes or variations derived therefrom are still within the protection scope of the present application.

Claims

1. A combined device for flotation and smelting of lead and zinc from difficult and complex ores, characterized by: It comprises a flotation system (7) and an oxygen pressure acid leaching system; the oxygen pressure acid leaching system comprises a stage I oxygen pressure acid leaching system (1), a stage II oxygen pressure acid leaching system (2), a desulfurization reactor (3), a lead selection reaction tank (4), a purification tank (5) and a zinc electrolysis tank (6); The stage I oxygen pressure acid leaching system (1) comprises a ball mill (11), a first slurry mixing tank (12), a stage I oxygen pressure autoclave (13), a stage I flash tank (14), a stage I intermediate tank (15) and a stage I filter press (16) connected in sequence; The stage II oxygen pressure acid leaching system (2) comprises a second slurry conditioning tank (21), a stage II oxygen pressure autoclave (22), a stage II flash tank (23), a stage II intermediate tank (24) and a stage II filter press (25) connected in sequence; The liquid outlet (161) of the filter press of stage I is connected to the purification tank (5), and the slag outlet (162) of the filter press of stage I is connected to the second slurry mixing tank (21); the liquid outlet (61) of the zinc electrolytic cell is connected to the second slurry mixing tank (21), the liquid outlet (251) of the filter press of stage II is connected to the first slurry mixing tank (12), the slag outlet (252) of the filter press of stage II is connected to the desulfurization reactor (3), and the slag outlet (332) of the desulfurization reactor is connected to the lead selection reaction tank (4); The flotation system (7) includes a plurality of flotators (8) connected in sequence, wherein the flotators (8) include a base (81), an overflow trough (82), a flotation motor (83) and an impeller (84); a flotation zone (85) is formed between the base (81) and the overflow trough (82); the lower end of the flotation zone (85) is connected to an air charging pipe (851) outwardly, and the other end of the air charging pipe (851) is connected to an air charging pump (852); the flotation motor (83) is arranged at the center of the base (81) and is connected to the impeller (84) via a rotating shaft extending from the base (81); the overflow trough (82) is arranged around the top of the flotation zone (85); A concentrate transmission pipe (86) is provided below the overflow trough (82) and is connected to the flotation zone (85) of the flotation device (8) at the next stage; A middling inlet (87) and a tailings outlet (88) are provided on both sides of the lower portion of the flotation zone (85), and the tailings outlet (88) of the flotation device (8) at the next level is connected to the middling inlet (87) of the flotation device (8) at the previous level via a tailings transmission pipe (89), and a peristaltic pump (891) is installed in the tailings transmission pipe (89); The desulfurization reactor (3) comprises a reaction tank (31), a tank cover (32), a solid-liquid separator (33), a rotary cover-opening motor (34) and a dual-motor controller (35); The top of the reaction tank (31) is connected to a tank cover (32), which is rotatably connected in parallel with the reaction tank (31), and a rotating shaft (36) is fixed to the edge of the upper port of the reaction tank (31); The lower end of the reaction tank (31) is in an inverted cone shape, and the bottom is connected to the solid-liquid separator (33). The solid-liquid separator (33) and the reaction tank (31) are rotatably connected in parallel, and the rotating shaft (36) is fixed to the edge of the lower end of the reaction tank (31); the rotating shaft (36) of the tank cover (32) and the solid-liquid separator (33) are each connected to a rotary cover opening motor (34), and the rotary cover opening motor (34) is connected to a dual motor controller (35). The dual motor controller (35) is arranged at the lower part of the tank body; The upper end of the solid-liquid separator (33) is a desulfurization reactor filter screen (331), and the lower end is a desulfurization reactor slag outlet (332). A desulfurization reactor liquid outlet pipe (333) is provided on the side of the lower portion of the solid-liquid separator (33).

2. The combined device for flotation and smelting of lead and zinc from refractory and complex ores according to claim 1, characterized in that: The flotation system (7) includes a roughing flotation cell (71), wherein both sides of the roughing flotation cell (71) are connected to a first fine flotation cell (72) and a first scavenging flotation cell (75), respectively; the first fine flotation cell (72) is connected to a second fine flotation cell (73) and a third fine flotation cell (74) in sequence; the first scavenging flotation cell (75) is connected to a second scavenging flotation cell (76) and a third scavenging flotation cell (77) in sequence; and a concentrate transmission pipe (86) of the third fine flotation cell (74) is connected to a ball mill (11).

3. The combined device for flotation and smelting of lead and zinc from refractory and complex ores according to claim 1, characterized in that: The lead selection reaction tank (4) is provided with a temperature control mechanism (41) and a solid-liquid separation tube (42). The temperature control mechanism (41) comprises a fixed seat (411), an electric heating wire (412), a cooling fan (413), a heating motor (414), a cooling motor (415) and a temperature control motor controller (416). The electric heating wires (412) are arranged on both sides of the fixed seat (411). The lead selection reaction tank (4) is placed in the fixed seat (411). The cooling fan (413) is placed The lower end of the electric heating wire (412) is connected to the heating motor (414), the cooling fan (413) is connected to the cooling motor (415), the heating motor (414) and the cooling motor (415) are arranged on the fixing seat (411) and connected to the temperature control motor controller (416), and the temperature control motor controller (416) is arranged at the lower part of the outer side of the fixing seat (411); the solid-liquid separation tube (42) is arranged at the front side of the lower end of the lead selection reaction tank (4).

4. The combined device for flotation and smelting of lead and zinc from refractory and complex ores according to claim 1, characterized in that The first slurry mixing tank (12) and the second slurry mixing tank (21) are provided with temperature control devices, which can adjust the reaction temperature in the tank.

5. A method for flotation-smelting lead and zinc from difficult-to-select complex ores, characterized by: The combined device for flotation and smelting lead and zinc from refractory and complex ores as described in any one of claims 1 to 4 specifically comprises the following steps: A. Grinding: Grind the raw ore to a fineness of -200 mesh, which accounts for 70-80% of the ore pulp, and transport it to the flotation system (7); B. Flotation of lead and zinc: Add the following reagents respectively and perform flotation according to the steps: ① Roughing: add 15kg / t of lime, 200g / t of ethyl thiocyanate, 200g / t of sodium isobutyl xanthate or potassium isobutyl xanthate, 300g / t of sodium hexametaphosphate, and 70g / t of pine oil; start the flotation motor (83), transport the concentrate foam from the overflow tank of the roughing flotation cell (71) to the first flotation cell (72), and transport the tailings to the first flotation cell (75) through the tailings transmission pipe (89); ②Scan and select: During the scavenging I process, 3500 g / t of lime, 200 g / t of ethyl thiocyanate, 200 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate, 300 g / t of sodium hexametaphosphate, 300 g / t of copper sulfate, and 50 g / t of pine oil are added; the flotation motor (83) is started to transport the concentrate foam from the overflow tank (82) of the scavenging I flotation cell (75) back to the roughing flotation cell (71), and the tailings are transported to the scavenging II flotation cell (76) through the tailings transmission pipe (89); During the scavenging II process, 2000 g / t of lime, 100 g / t of ethyl thiocyanate, 100 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate, 150 g / t of sodium hexametaphosphate, 150 g / t of copper sulfate, and 30 g / t of pine oil are added; the flotation motor (83) is started to transport the concentrate foam from the overflow tank (82) of the scavenging second flotation device (76) back to the scavenging first flotation device (75), and the tailings are transported to the scavenging third flotation device (77) through the tailings transmission pipe (89); During the scavenging III process, 50 g / t of ethyl thiocyanate, 50 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate, 50 g / t of sodium hexametaphosphate, 50 g / t of copper sulfate, and 10 g / t of pine oil are added; the flotation motor (83) is started to transport the concentrate foam from the overflow tank (82) of the scavenging III flotation device (77) back to the scavenging II flotation device (76), and the tailings are collected and processed through the tailings transmission pipe (89); ③Selected: In the concentration I process, 1500 g / t of lime, 200 g / t of ethyl thiocyanate, and 200 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate are added; the flotation motor (83) is started to transport the concentrate foam from the overflow tank (82) of the concentration first flotation device (72) to the concentration second flotation device (73), and the tailings are transported back to the roughing flotation device (71) through the tailings transmission pipe (89); During the concentration II process, 1000 g / t of lime, 150 g / t of ethyl thiocyanate, and 150 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate are added; the flotation motor (83) is started to transport the concentrate foam from the overflow tank (82) of the concentration II flotation device (73) to the concentration III flotation device (74), and the tailings are transported back to the concentration I flotation device (72) via the tailings transmission pipe (89); In the concentration III process, 500 g / t of lime, 60 g / t of ethyl thiocyanate, and 60 g / t of sodium isobutyl xanthate or potassium isobutyl xanthate are added; the flotation motor (83) is started, and the mixed lead-zinc concentrate produced by the overflow tank (82) of the concentration third flotation device (74) is transported to the oxygen pressure acid leaching system (1) of the first stage through the concentrate transmission pipe (86); the tailings are transported back to the concentration second flotation device (73) through the tailings transmission pipe (89); C. Pressure oxidation acid leaching ① Slurry preparation: The mixed lead-zinc concentrate and a sulfuric acid solution with a mass concentration of 140 g / L are mixed at a solid-liquid mass ratio of 1:(3-5) and transported into a ball mill (11) to be ground to a particle size of -200-100 mesh, and then fed into a first slurry mixing tank (12), and sodium lignin sulfonate is added to obtain a slurry mixing liquid, wherein the amount of sodium lignin sulfonate added is 0.2-0.5 wt% of the mixed lead-zinc concentrate; ②Ⅰ stage pressure oxidation acid leaching: After being preheated to 80-120°C in the first slurry mixing tank (12), the mixture is added to the first stage oxygen autoclave (13), heated to 150-180°C, and oxygen is introduced to increase the pressure to 0.6-2.0 MPa, and the first stage pressure oxidation acid leaching is carried out, with a reaction time of 60-180 min. The mixture then passes through the first stage flash tank (14) for flash evaporation and pressure reduction, and enters the first stage intermediate tank (15). The mixture then passes through the first stage filter press (16) for solid-liquid separation, and obtains the first stage leaching solution and the first stage leaching residue. ③Ⅱ stage pressure oxidation acid leaching The waste electrolyte after electrolysis of the leaching residue of stage I and the leaching solution of stage I is added to the second slurry mixing tank (21) for slurry mixing according to the liquid-solid mass ratio (4-10):1, and then preheated to 80-120°C and added to the oxygen autoclave (22) of stage II, and oxygen is introduced and pressurized to 0.4-1.2 MPa, and stage II pressure oxidation acid leaching is carried out, and the reaction time is 100-360 minutes. Then, the stage II flash tank (23) is flashed and depressurized, and then the stage II intermediate tank (24) is entered, and then the stage II leaching solution and stage II leaching residue are obtained after solid-liquid separation by the stage II filter press (25). The stage II leaching solution returns to the first slurry mixing tank (12) and continues to participate in the pressure oxidation acid leaching of stage I. ④Ⅱ stage leaching residue treatment: Desulfurization: The leaching residue from stage II is fed into the desulfurization reactor (3), and a 4 mol / L ammonium sulfide solution is added at a liquid-to-solid volume mass ratio of 11 mL / g. The reaction is maintained at room temperature for 40 to 50 minutes, and the sulfur leaching rate can reach 98%. The filtrate is separated and subjected to thermal decomposition to obtain sulfur. Lead selection: The filter residue from the desulfurization process enters the lead selection reaction tank (4), and a nitric acid solution with a concentration of 7 mol / L is added according to a liquid-solid ratio of 4:

1. The leaching time is 50 minutes and the leaching temperature is 80°C. After cooling, the filtrate is filtered and washed. Sulfuric acid is added to the filtrate to form lead sulfate precipitation. After filtering, the lead is separated and the lead sulfate is converted into lead acetate, which is then replaced with lead; ⑤Ⅰ stage leaching solution treatment: The leachate from stage I is first discharged into the purification tank (5) and then into the zinc electrolytic tank (6). After separation, purification and electrolysis, waste electrolyte and electrolytic zinc are obtained respectively. The waste electrolyte returns to the intermediate tank (15) of stage I and continues to pass through the filter press (16) of stage I for solid-liquid separation. The electrolytic zinc is then melted and cast to obtain zinc ingots.

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