CN103409622A - Method for individually processing high-iron zinc sulfide concentrate - Google Patents
Method for individually processing high-iron zinc sulfide concentrate Download PDFInfo
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Abstract
The invention belongs to the technical field of hydrometallurgy, and particularly relates to a method for individually processing high-iron zinc sulfide concentrate. The method comprises a step of subjecting the high-iron zinc sulfide concentrate to calcination in a fluidized bed combustion boiler to obtain zinc calcine; a step of subjecting the zinc calcine to neutral leaching to produce neutral leaching solution and neutral leaching residue; a step of, after the neutral leaching residue and the high-iron zinc sulfide concentrate are mixed, successively performing reduction leaching and oxidation leaching, and circulating oxidation leaching solution to the reduction leaching to produce reduction leaching solution and silver-rich sulfur residue; a step of replacing the reduction leaching solution by using iron powder to precipitate copper and to produce copper-rich slag and solution after copper precipitation; a step of subjecting the solution after copper precipitation to pre-neutralization by using the zinc calcine, and then replacing by using zinc powder to precipitate indium and to produce indium-rich slag and solution after indium precipitation; and a step of bubbling oxygen into the solution after indium precipitation, heating and removing iron to obtain iron removal solution and hematite slag. The hematite slag can be utilized as a raw material for ironmaking. The method has strong pertinence, short technological process and high metal recovery yield, and the method is clean, efficient, energy-saving and environmental friendly. Separation and comprehensive utilization of zinc, indium, copper and iron are achieved.
Description
Technical field
The invention belongs to technical field of wet metallurgy, particularly relate to a kind of method of individual curing high-iron zinc sulfide concentrate.
Background technology
Marmatite is in becoming the ore deposit process, iron, indium replace the zinc atom in zink sulphide with isomorph, adopt the method for mechanical ore grinding and ore dressing to be difficult to make zinc, iron, indium to separate, causing the zinc ore concentrate iron content of ore dressing output, high to contain zinc low, and association has the metals such as the indium, copper of high level, be called high-iron zinc sulfide concentrate.The common zinc grade of this high-iron zinc sulfide concentrate is lower, and indium, iron level are high, and associate lead content is low, and association simultaneously has a considerable number of copper and tin.
In smelting process, because the iron in high-iron zinc sulfide concentrate and zinc are inlayed and coexisted, under the condition of roasting, inevitably produce a large amount of zinc ferrites, zinc ferrite is not destroyed at neutral leaching process, and with not dissolved zinc oxide, and most iron, indium, copper are stayed in neutral leaching residue.
In order to destroy zinc ferrite, in order to reclaim zinc and indium, usually need to adopt rotary kiln evaporation method or hot acid leaching-out technique to process.Wherein rotary kiln evaporation method can effectively destroy zinc ferrite, and output zinc oxide and Indium sesquioxide return to the zinc metallurgy flow process and reclaim the zinc indium.Yet the rotary kiln evaporation method energy consumption is high, the zinc indium recovery is low, and the output low concentration sulphur dioxide flue gas is difficult to administer, and has limited applying of the method.It is a kind of effective ways that destroy zinc ferrite that hot acid leaches, and zinc, iron, indium together enter in solution, and the iron major part is with Fe
3+Form exists, and causes the iron difficult separation and recycling.At present, from the method for separation of iron this solution, be mainly jarosite process and goethite process.While adopting the jarosite process separation of iron, indium and iron together enter Jarosite Residues, then adopt volatilization method from iron vitriol slag, reclaiming indium.Goethite process needs first by the Fe in solution
3+Be reduced to Fe
2+, by the neutralizing hydrolysis method, reclaim indium, obtain indium slag.After heavy indium, to continue to be neutralized to pH be 2.5～4.2 to liquid, under 85～90 ℃ of conditions, adopts the atmospheric oxidation deironing, obtains the pyrrhosiderite slag.Yet above-mentioned two kinds of hot acids leaching method for removing iron scum iron content is low, the quantity of slag is large, and the scum obtained can't utilize.Therefore, high-iron zinc sulfide concentrate is used usually used as batching for a long time, fails to be developed preferably.
In order to solve the difficult problem of high-iron zinc sulfide concentrate comprehensive utilization, numerous scientific workers have carried out a large amount of explorations and technological improvement, and have developed some new Technologies.The patent that is 201010300159.7 as (1) Chinese patent application number discloses a kind of method that slag-free zinc hydrometallurgy of zinc concentrate is carried indium and produced ferric oxide.The technological line adopted be " fluosolids roasting---the neutral leaching---low Ore Leaching---high Ore Leaching---prereduction neutralization---displacement copper removal---neutralize sink indium---hydrothermal method is sunk iron ".(2) to be that 201110286157.1 patent discloses a kind of from the high Indium-Zinc concentrate of high ferro, extracting the zinc indium and reclaiming the method for iron for number of patent application.The technical process of adopting be " fluidized bed roasting-neutrality leaches---hot acid reduces leaching---zinc oxide pre-neutralization---zinc dust precipitation sink indium---hematite process sinks iron ".These methods all effectively raise the rate of recovery of zinc, indium, and have fully utilized the iron in mineral, but in mineral, the copper of association and tin do not obtain efficient recovery.
Summary of the invention
The objective of the invention is specially for high-iron zinc sulfide concentrate, solve high efficiency separation and the comprehensive problem reclaimed of zinc, iron, indium, copper in smelting process, provide a kind of high-iron zinc sulfide concentrate to extract the processing method of zinc indium copper and iron resources utilization, improve metal recovery rate and comprehensive utilization of resources rate, environmental contamination reduction.
Realize that the technical scheme that the object of the invention is taked is: high-iron zinc sulfide concentrate is carried out to roasting in fluidizing furnace, obtain zinc baking sand; Zinc baking sand carries out neutrality and leaches, the neutral leach liquor of output and neutral leaching residue; Neutral leaching residue with after high-iron zinc sulfide concentrate mixes successively reduction leach and Oxidation Leaching, Oxidation Leaching liquid is circulated to reduction and leaches, output reduction leach liquor and Fu Yin sulphur slag; The reduction leach liquor is with the heavy copper of iron replacement, liquid after the rich copper ashes of output and heavy copper; After heavy copper liquid after the zinc baking sand pre-neutralization with the heavy indium of zinc dust precipitation, liquid after output indium slag and heavy indium; After heavy indium, liquid passes into oxygen and heats deironing, obtains liquid and rhombohedral iron ore slag after deironing, and the rhombohedral iron ore slag is as the iron-smelting raw material recycling.
Described high-iron zinc sulfide concentrate be contain zinc 35～46wt%, iron content 16～25wt%, contain indium 300～1200g/t, the high-iron zinc sulfide concentrate of cupric 0.1～1.5wt%, stanniferous 0.3～1.0wt%.
Order of the present invention technical scheme more specifically is:
1. after the high-iron zinc sulfide concentrate roasting, should make sulphur content≤2wt% in zinc baking sand;
2. the neutral leaching agent that leaches employing of zinc baking sand is the solution of the sulfur acid 80～100g/L of liquid preparation after zinc electrolysis waste solution and deironing, the neutral operation that leaches is carried out according to the zinc hydrometallurgy routine techniques, and neutral leach liquor enters the zinc hydrometallurgy operation and purifies the electric zinc of electrodeposition production;
3. to leach the mass ratio that the neutral leaching residue of operation mixes with high-iron zinc sulfide concentrate be 1:0.15～1:0.35 in reduction, leaching agent is the solution that zinc electrolysis waste solution and Oxidation Leaching liquid add the sulfur acid 180～220g/L of part bright sulfur acid preparation, 85 ℃～95 ℃ of temperature of reaction, in 2～4 hours reaction times, reaction proceeds to reduction leach liquor sulfur acid 15～30g/L, Fe
3+Be less than 3 g/L;
4. the Oxidation Leaching operation is added the part neutral leaching residue, mass ratio 1:0.65～1:1 that the reduction leached mud mixes with neutral leaching residue, leaching agent is the solution that the zinc electrolysis waste solution adds the sulfur acid 180～220g/L of part bright sulfur acid preparation, 85 ℃～95 ℃ of temperature of reaction, 2～4 hours reaction times, reaction obtains rich silver-colored sulphur slag and contains that zinc is less than 4wt%, sulfur-bearing is greater than 32wt%, and argentiferous is greater than 500g/t;
5. the heavy copper process of displacement adopts the multi-stage countercurrent displacement, and reaction proceeds to the rich copper ashes cupric >=50wt% obtained after heavy copper;
6. the heavy indium operation of the displacement method that adopts one section zinc baking sand pre-neutralization to add one section zinc dust precipitation is carried out, the liquid after copper that namely first will sink is heated to 70～85 ℃, in liquid after heavy copper, add zinc baking sand 30～60 g/L solution again, liquid pH value to 1.5～2.0 after the heavy copper of adjusting, reacted 20～40 minutes, carry out liquid-solid separation, liquid in output and after slag and pre-neutralization, in and slag return former leaching; Then under 80～90 ℃, add zinc powder 6～10g/L solution in liquid after pre-neutralization, reacted 30～60 minutes, obtain liquid after heavy indium and contain the indium slag of indium 0.8～2.5wt%;
7. the pressure that the iron removal step of liquid should pass into oxygen after heavy indium is 1.6～2.5MPa, and Heating temperature is 180 ℃～200 ℃.
The maturing temperature of described high-iron zinc sulfide concentrate in fluidizing furnace is 950～1100 ℃, and under this roasting condition, in high-iron zinc sulfide concentrate, most metallic sulfides can be oxidized to metal oxide, and sulphur is oxidized to sulfurous gas, for relieving haperacidity; The not oxidized sulfide of minute quantity enters into zinc baking sand, obtains the zinc baking sand of sulfur-bearing hardly.
The object of the invention technical scheme also comprises: the leaching liquid-solid ratio 4～8mL/g of operation, the leaching liquid-solid ratio 4～8mL/g of Oxidation Leaching operation are leached in reduction; The heavy copper process of displacement of reduction leach liquor first is heated to 75～85 ℃ for first reducing leach liquor, then adds iron powder 3～6g/L solution to reducing in leach liquor, reacts 20～40 minutes; After heavy indium, the reaction times of the iron removal step of liquid is 2～4 hours, and after reaching deironing, the liquid iron content is less than 2g/L, contains zinc 90～110g/L, and rhombohedral iron ore slag iron content 55%～65wt%, contain zinc≤0.5%, contain arsenic≤0.01%, sulfur-bearing≤6%.
Beneficial effect of the present invention is as follows:
(1) technique is with strong points.The present invention is directed to high efficiency separation and the comprehensive problem reclaimed of zinc, iron, indium, copper in the high-iron zinc sulfide concentrate smelting process, proposed a kind of method of individual curing high-iron zinc sulfide concentrate;
(2) comprehensive utilization of resources rate is high.The present invention has adopted reduction leaching and two sections counterflow leaching techniques of Oxidation Leaching of neutral leaching residue, realized that total leaching yield of zinc is more than 99wt%, total leaching yield of indium is more than 95wt%, total leaching yield of copper is more than 98wt%, the finishing slag rate is less than 13wt%, finishing slag zinc content is less than 4wt%, indium content is less than 100g/t, and the metals such as lead, silver, tin concentration ratio in leaching finishing slag is high, is conducive to the recovery of plumbous silver-colored tin.Whole technique zinc recovery is greater than 95%, and indium recovery is greater than 85%, and copper recovery is greater than 90%, and iron recovery is greater than 80%;
(3) smelting process clean and effective.The present invention has saved traditional Rotary Kiln technique, has reduced energy consumption, has avoided smoke pollution, has optimized technical process; Adopt the deironing of oxygen setting-out thermal technology, realize that zinc-iron separates, obtain the rhombohedral iron ore slag, improved the rate of recovery of zinc indium copper, realized the recycling of iron, effectively solve simultaneously the acid balance of high-iron zinc sulfide concentrate smelting process.
The accompanying drawing explanation
Fig. 1 is process flow sheet of the present invention.
Specific embodiment
The high-iron zinc sulfide concentrate that certain factory provides of take is raw material, and (wt%) is as follows for its composition: Zn:44.3, Fe16.79, S:32.55, SiO
2: 4.45, Cu:0.66, Pb:0.06, In:0.0395.Through high-temperature roasting, obtain the zinc baking sand of following composition (wt%):: Zn:51.78, Fe18.58, S:0.73, SiO
2: 3.19, Cu:0.81, Pb:0.07, In:0.0458.
Embodiment 1:
1, the waste electrolyte 214L of liquid 376L, sulfur acid 165g/L, zinc 37.5g/L after the deironing of 100kg zinc baking sand and sulfur acid 45g/L, zinc 98g/L is mixed and carry out neutrality and leach, obtain containing the neutral leach liquor of zinc 138g/L, neutral leach liquor send purifying electrolysis; Output neutral leaching residue 65kg, its composition following (wt%): Zn:25.94, Fe:26.58, In:0.051, Cu:1.00, Pb:0.099.
2, getting above-mentioned neutral leaching residue 51.2kg mixes with the 15.3kg high-iron zinc sulfide concentrate, with Oxidation Leaching liquid 173L, sulfur acid 165g/L, the waste electrolyte 182L of zinc 37.5g/L, the concentration of sulfur acid 30g/L, zinc 46g/L, be that 98% vitriol oil 23L and 21L water mix and reduces leaching, 90 ℃ of extraction temperatures, extraction time 3h, obtain reducing leach liquor 385L, output reduction leached mud 14.6kg.Reduction leach liquor composition (g/L) is as follows: Zn:94.14, TFe:36.65, Fe
2+: 33.06, In:0.15, Cu:1.74, H
2SO
4: 20.
3,14.6kg being reduced to leached mud mixes with the 13.8kg neutral leaching residue, with waste electrolyte 147.7L, the concentration of sulfur acid 165g/L, zinc 37.5g/L, be that 98% vitriol oil 6.8L and 15.9L water mix and carries out Oxidation Leaching, 90 ℃ of extraction temperatures, extraction time 3h, obtain Oxidation Leaching liquid 164L, the rich silver-colored sulphur slag 12.4kg of output.Reduction leached mud composition (wt%) is as follows: Zn:3.64, Fe:6.12, In:0.0083, Cu:0.14, S
0: 28.7, Ag:0.0512.
4, get reduction leach liquor 100L, be heated 85 ℃ of backward its and add iron powder 6g/L to carry out replacement(metathesis)reaction, reaction times 20min, the rich copper ashes 235g of output cupric 62%.
5, to adding zinc baking sand 30g/L in liquid after the heavy copper of 100L, carry out pre-neutralization, 70 ℃ of neutral temperatures, reaction times 40min, obtain terminal pH1.58, contain liquid after the pre-neutralization of zinc 99g/L, in and slag return to reduction and leach.
6, in liquid after the 50L pre-neutralization, add zinc powder 8g/L solution, react 40min under 90 ℃, obtain liquid after heavy indium and contain the indium slag of indium 1.64wt%.
7, liquid pump after the heavy indium of 35L is entered in autoclave, be heated to 180 ℃, then pass into oxygen, control stagnation pressure 1.6Mpa, reaction 3h; Obtain iron content 1.45g/L, contain liquid and iron content 61% after the deironing of zinc 98g/L, contain zinc 0.33%, contain arsenic 0.004%, the rhombohedral iron ore slag of sulfur-bearing 5.68%; After deironing, liquid returns to neutral the leaching.
Embodiment 2:
1, the waste electrolyte 212L of liquid 378L, sulfur acid 165g/L, zinc 37.5g/L after the deironing of 100kg zinc baking sand and sulfur acid 48g/L, zinc 105g/L is mixed and carry out neutrality and leach, obtain containing the neutral leach liquor of zinc 141g/L, neutral leach liquor send purifying electrolysis; Output neutral leaching residue 62kg, composition following (wt%): Zn:25.94, Fe:26.58, In:0.051, Cu:1.00, Pb:0.099.
2, getting above-mentioned neutral leaching residue 52kg mixes with the 11kg high-iron zinc sulfide concentrate, with Oxidation Leaching liquid 130L, sulfur acid 165g/L, the waste electrolyte 186L of zinc 37.5g/L, the concentration of sulfur acid 28g/L, zinc 42g/L, be that 98% vitriol oil 25L and 33L water mix and reduces leaching, 95 ℃ of extraction temperatures, extraction time 2h, obtain reducing leach liquor 358L, output reduction leached mud 12.3kg.Reduction leach liquor composition (g/L) is as follows: Zn:85.5, Tfe:37.82, Fe
2+: 33.32, In:0.16, Cu:1.93, H
2SO
4: 26.
3,12.3kg being reduced to leached mud mixes with the 10kg neutral leaching residue, with waste electrolyte 107L, the concentration of sulfur acid 165g/L, zinc 37.5g/L, be that 98% vitriol oil 4.5L and 11L water mix and carries out Oxidation Leaching, 95 ℃ of extraction temperatures, extraction time 2h, obtain Oxidation Leaching liquid 118L, the rich silver-colored sulphur slag 8.2kg of output.Reduction leached mud composition (wt%) is as follows: Zn:3.26, Fe:5.87, In:0.0079, Cu:0.13, S
0: 26.7, Ag:0.0536.
4, get reduction leach liquor 100L, be heated 80 ℃ of backward its and add iron powder 3g/L to carry out replacement(metathesis)reaction, reaction times 30min, the rich copper ashes 291g of output cupric 66.4%.
5, to adding zinc baking sand 48g/L in liquid after the heavy copper of 100L, carry out pre-neutralization, 75 ℃ of neutral temperatures, reaction times 30min, obtain terminal pH1.72, contain liquid after the pre-neutralization of zinc 103g/L, in and slag return to reduction and leach.
6, in liquid after the 50L pre-neutralization, add zinc powder 6g/L solution, react 30min under 80 ℃, obtain liquid after heavy indium and contain the indium slag of indium 2.50wt%.
7, liquid pump after the heavy indium of 35L is entered in autoclave, be heated to 190 ℃, then pass into oxygen, control stagnation pressure 2.0Mpa, reaction 4h; Obtain iron content 1.26g/L, contain liquid and iron content 62.7% after the deironing of zinc 100g/L, contain zinc 0.21%, contain arsenic 0.005%, the rhombohedral iron ore slag of sulfur-bearing 4.57%; After deironing, liquid returns to neutral the leaching.
Embodiment 3:
1, the waste electrolyte 215L of liquid 375L, sulfur acid 165g/L, zinc 37.5g/L after the deironing of 100kg zinc baking sand and sulfur acid 46g/L, zinc 103g/L is mixed and carry out neutrality and leach, obtain containing the neutral leach liquor of zinc 138g/L, neutral leach liquor send purifying electrolysis; Output neutral leaching residue 64kg, composition following (wt%): Zn:25.94, Fe:26.58, In:0.051, Cu:1.00, Pb:0.099.
2, getting above-mentioned neutral leaching residue 53kg mixes with the 12.5kg high-iron zinc sulfide concentrate, with Oxidation Leaching liquid 146L, sulfur acid 165g/L, the waste electrolyte 180L of zinc 37.5g/L, the concentration of sulfur acid 30g/L, zinc 45g/L, be that 98% vitriol oil 24L and 27L water mix and reduces leaching, 85 ℃ of extraction temperatures, extraction time 4h, obtain reducing leach liquor 372L, output reduction leached mud 13.2kg.Reduction leach liquor composition (g/L) is as follows: Zn:90.2, Tfe:37.23, Fe
2+: 34.26, In:0.16, Cu:1.82, H
2SO
4: 24.
3,13.2kg being reduced to leached mud mixes with the 11kg neutral leaching residue, with waste electrolyte 144L, the concentration of sulfur acid 165g/L, zinc 37.5g/L, be that 98% vitriol oil 5.26L and 1.4L water mix and carries out Oxidation Leaching, 85 ℃ of extraction temperatures, extraction time 4h, obtain Oxidation Leaching liquid 146L, the rich silver-colored sulphur slag 9.8kg of output.Reduction leached mud composition (wt%) is as follows: Zn:2.98, Fe:5.45, In:0.0073, Cu:0.12, S
0: 30.2, Ag:0.0563.
4, get reduction leach liquor 100L, be heated 75 ℃ of backward its and add iron powder 4.5g/L to carry out replacement(metathesis)reaction, reaction times 40min, the rich copper ashes 285g of output cupric 63.7%.
5, to adding zinc baking sand 60g/L in liquid after the heavy copper of 100L, carry out pre-neutralization, 85 ℃ of neutral temperatures, reaction times 20min, obtain terminal pH1.92, contain liquid after the pre-neutralization of zinc 102g/L, in and slag return to reduction and leach.
6, in liquid after the 50L pre-neutralization, add zinc powder 10g/L solution, react 60min under 85 ℃, obtain liquid after heavy indium and contain the indium slag of indium 0.81wt%.
7, liquid pump after the heavy indium of 35L is entered in autoclave, be heated to 200 ℃, then pass into oxygen, control stagnation pressure 2.5Mpa, reaction 2h; Obtain iron content 1.98g/L, contain liquid and iron content 57% after the deironing of zinc 103g/L, contain zinc 0.42%, contain arsenic 0.007%, the rhombohedral iron ore slag of sulfur-bearing 5.06%; After deironing, liquid returns to neutral the leaching.
Claims (5)
1. the method for an individual curing high-iron zinc sulfide concentrate, is characterized in that: high-iron zinc sulfide concentrate is carried out to roasting in fluidizing furnace, obtain zinc baking sand; Zinc baking sand carries out neutrality and leaches, the neutral leach liquor of output and neutral leaching residue; Neutral leaching residue with after high-iron zinc sulfide concentrate mixes successively reduction leach and Oxidation Leaching, Oxidation Leaching liquid is circulated to reduction and leaches, output reduction leach liquor and Fu Yin sulphur slag; The reduction leach liquor is with the heavy copper of iron replacement, liquid after the rich copper ashes of output and heavy copper; After heavy copper liquid after the zinc baking sand pre-neutralization with the heavy indium of zinc dust precipitation, liquid after output indium slag and heavy indium; After heavy indium, liquid passes into oxygen and heats deironing, obtains liquid and rhombohedral iron ore slag after deironing.
2. the method for individual curing high-iron zinc sulfide concentrate according to claim 1 is characterized in that: described high-iron zinc sulfide concentrate for containing zinc 35～46wt%, iron content 16～25wt%, contain indium 300～1200g/t, cupric 0.1～1.5wt%, stanniferous 0.3～1.0wt%.
3. the method for individual curing high-iron zinc sulfide concentrate according to claim 1 is characterized in that:
1. after the high-iron zinc sulfide concentrate roasting, should make sulphur content≤2wt% in zinc baking sand;
2. the neutral leaching agent that leaches employing of zinc baking sand is the solution of the sulfur acid 80～100g/L of liquid preparation after zinc electrolysis waste solution and deironing, the neutral operation that leaches is carried out according to the zinc hydrometallurgy routine techniques, and neutral leach liquor enters the zinc hydrometallurgy operation and purifies the electric zinc of electrodeposition production;
3. to leach the mass ratio that the neutral leaching residue of operation mixes with high-iron zinc sulfide concentrate be 1:0.15～1:0.35 in reduction, leaching agent is the solution that zinc electrolysis waste solution and Oxidation Leaching liquid add the sulfur acid 180～220g/L of part bright sulfur acid preparation, 85 ℃～95 ℃ of temperature of reaction, in 2～4 hours reaction times, reaction proceeds to reduction leach liquor sulfur acid 15～30g/L, Fe
3+Be less than 3 g/L;
4. the Oxidation Leaching operation is added the part neutral leaching residue, mass ratio 1:0.65～1:1 that the reduction leached mud mixes with neutral leaching residue, leaching agent is the solution that the zinc electrolysis waste solution adds the sulfur acid 180～220g/L of part bright sulfur acid preparation, 85 ℃～95 ℃ of temperature of reaction, 2～4 hours reaction times, reaction obtains rich silver-colored sulphur slag and contains that zinc is less than 4wt%, sulfur-bearing is greater than 32wt%, and argentiferous is greater than 500g/t;
5. the heavy copper process of displacement adopts the multi-stage countercurrent displacement, and reaction proceeds to the rich copper ashes cupric >=50wt% obtained after heavy copper;
6. the heavy indium operation of the displacement method that adopts one section zinc baking sand pre-neutralization to add one section zinc dust precipitation is carried out, the liquid after copper that namely first will sink is heated to 70～85 ℃, in liquid after heavy copper, add zinc baking sand 30～60 g/L solution again, liquid pH value to 1.5～2.0 after the heavy copper of adjusting, reacted 20～40 minutes, carry out liquid-solid separation, liquid in output and after slag and pre-neutralization, in and slag return former leaching; Then under 80～90 ℃, add zinc powder 6～10g/L solution in liquid after pre-neutralization, reacted 30～60 minutes, obtain liquid after heavy indium and contain the indium slag of indium 0.8～2.5wt%;
7. the pressure that the iron removal step of liquid should pass into oxygen after heavy indium is 1.6～2.5MPa, and Heating temperature is 180 ℃～200 ℃.
4. according to the method for the described individual curing high-iron zinc sulfide concentrate of claim 2 or 3, it is characterized in that: the leaching liquid-solid ratio 4～8mL/g of operation, the leaching liquid-solid ratio 4～8mL/g of Oxidation Leaching operation are leached in reduction; The heavy copper process of displacement of reduction leach liquor first is heated to 75～85 ℃ for first reducing leach liquor, then adds iron powder 3～6g/L solution to reducing in leach liquor, reacts 20～40 minutes; After heavy indium, the reaction times of the iron removal step of liquid is 2～4 hours, and after reaching deironing, the liquid iron content is less than 2g/L, contains zinc 90～110g/L, and rhombohedral iron ore slag iron content 55%～65wt%, contain zinc≤0.5%, contain arsenic≤0.01%, sulfur-bearing≤6%.
5. according to the method for the described individual curing high-iron zinc sulfide concentrate of claim 2 or 3, it is characterized in that: the maturing temperature of high-iron zinc sulfide concentrate in fluidizing furnace is 950～1100 ℃, the sulfurous gas produced is for relieving haperacidity, and not oxidized sulfide enters into zinc baking sand.
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| CN103789544A (en) * | 2014-02-13 | 2014-05-14 | 昆明理工大学科技产业经营管理有限公司 | Synergistic leaching-copper arsenate removing method for leaching residues in high-iron zinc calcine and high-iron zinc sulfide concentrate |
| CN104004923A (en) * | 2014-06-11 | 2014-08-27 | 长沙有色冶金设计研究院有限公司 | Method for extracting zinc by combination of roasting leaching and direct leaching of zinc sulfide concentrate |
| CN104120253A (en) * | 2014-07-28 | 2014-10-29 | 蒙自矿冶有限责任公司 | Leaching method of complex zinc calcined ores |
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| CN104988325A (en) * | 2015-06-17 | 2015-10-21 | 广东省工业技术研究院(广州有色金属研究院) | Method for separating valuable metals from wet-process zinc smelting waste residues |
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| CN105838879B (en) * | 2016-03-29 | 2019-02-12 | 云南华联锌铟股份有限公司 | From the method and apparatus for removing removing calcium and magnesium after the heavy indium of zinc abstraction in liquid |
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