WO2019071787A1 - 一种由含镍冶炼熔渣回收有价组分的方法 - Google Patents

一种由含镍冶炼熔渣回收有价组分的方法 Download PDF

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Publication number
WO2019071787A1
WO2019071787A1 PCT/CN2017/115642 CN2017115642W WO2019071787A1 WO 2019071787 A1 WO2019071787 A1 WO 2019071787A1 CN 2017115642 W CN2017115642 W CN 2017115642W WO 2019071787 A1 WO2019071787 A1 WO 2019071787A1
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slag
iron
smelting
copper
reaction
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PCT/CN2017/115642
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English (en)
French (fr)
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张力
张武
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东北大学
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Publication of WO2019071787A1 publication Critical patent/WO2019071787A1/zh

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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/04—Working-up slag
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00—Obtaining noble metals
    • C22B11/02—Obtaining noble metals by dry processes
    • C22B11/021—Recovery of noble metals from waste materials
    • C22B11/023—Recovery of noble metals from waste materials from pyrometallurgical residues, e.g. from ashes, dross, flue dust, mud, skim, slag, sludge
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B13/00—Obtaining lead
    • C22B13/02—Obtaining lead by dry processes
    • C22B13/025—Recovery from waste materials
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B15/00—Obtaining copper
    • C22B15/0026—Pyrometallurgy
    • C22B15/0056—Scrap treating
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00—Obtaining zinc or zinc oxide
    • C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/02—Obtaining nickel or cobalt by dry processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/001—Dry processes
    • C22B7/004—Dry processes separating two or more metals by melting out (liquation), i.e. heating above the temperature of the lower melting metal component(s); by fractional crystallisation (controlled freezing)
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the invention belongs to the technical field of slag metallurgy, and in particular relates to a method for recovering valuable components from nickel-containing smelting slag.
  • the nickel smelting slag includes the nickel smelting slag produced by the “smelting smelting” process, the depleted slag after the “copper ice nickel blowing” process, and the top blowing smelting.
  • Nickel settled slag China Jinchuan Company uses “flash smelting” and “top blowing smelting” to carry out nickel concentrate smelting, which produces 1 million tons of nickel smelting slag per year. The accumulated accumulated storage capacity is over 10 million tons.
  • the valuable components such as copper, iron, nickel, cobalt, zinc, lead, gold and silver in the nickel smelting slag are stacked for a long time, which wastes resources and pollutes the environment.
  • the iron content in the nickel smelting slag is as high as 50wt%, far exceeding the current 26wt% recoverable grade, and the slag contains up to 0.2wt% copper, of which the iron component is mainly present in the fayalite phase.
  • the research and utilization of nickel smelting slag mainly focuses on direct reduction of iron or molten iron. During the reduction process, the reduction of copper, nickel and cobalt components into molten iron is not conducive to the subsequent steelmaking process.
  • Nickel smelting slag is not only an important secondary resource, but also contains a large amount of metallurgical fluxes such as SiO 2 , CaO, MgO, Al 2 O 3 , etc., which has strong chemical reaction activity and is a slag system with excellent physical and chemical properties.
  • the material is a mature metallurgical slag system. How to effectively use nickel smelting slag as the reaction slag system, and realize the simultaneous separation technology of copper and iron in nickel smelting slag is currently required by the market.
  • the present invention is a method for recovering valuable components from a nickel-containing smelting slag.
  • the method of the invention is a new slag metallurgy process, which not only reduces the copper content of the slag to make the slag contain copper ⁇ 0.1 wt%, but also realizes copper, nickel, cobalt, iron, gold, silver, lead, zinc, indium, antimony and sodium.
  • High-efficiency recovery of potassium and other components obtaining low-copper iron-containing materials (iron concentrate and pig iron), and solving the current problem of massive accumulation of slag, environmental pollution, and heavy metal pollution.
  • the main technical solutions adopted by the present invention include:
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • S1 slag mixing: adding nickel smelting slag to the smelting reaction device, adding calcium minerals and additives to form mixed slag, heating the mixed slag to a molten state as reaction slag, mixing uniformly, monitoring the reaction slag in real time By adjusting the mixed reaction slag to satisfy the condition a and the condition b at the same time, obtaining the slag after the reaction, or pouring the slag after the reaction into the heat preservation device;
  • condition a is controlling the temperature of the reaction slag to be 1100 to 1500 ° C;
  • the slag after the reaction in the step S1 is kept for 5 to 50 minutes, and the copper-rich nickel phase at the bottom, the iron-rich phase layer in the middle and lower portions, and the iron-containing silicate mineral phase layer in the middle and upper portions are obtained by sedimentation and separation.
  • the zinc component and the lead component are formed, the gold component and the silver component enter the copper-rich nickel phase, and the nickel-containing component and the cobalt component respectively enter the copper-rich nickel phase and the iron-rich phase; Recycling is carried out.
  • the method for regulating the condition a is:
  • the heating function of the reaction device itself is used, or the fuel and the preheated oxidizing air are added to the slag to make the temperature of the reaction slag reach the set temperature range, so that the temperature of the reaction slag reaches the set temperature range.
  • Reaction The temperature of the slag reaches 1100 ⁇ 1500 ° C;
  • reaction slag When the temperature of the reaction slag is >1500 ° C, one or more of nickel smelting slag, blast furnace slag, iron alloy slag, steel slag, metallurgical flux, iron-containing material or fluorine-containing material are added to the reaction slag, so that The temperature of the mixed slag reaches 1100 ⁇ 1500 ° C;
  • the method for regulating the condition b is:
  • the reaction device is a heat preservation device or a rotatable smelting reaction device or a smelting reaction device with slag or iron slag flowing out; wherein the heat preservation device is pourable Smelting reaction slag irrigation or insulation pit;
  • the rotatable smelting reaction device is a converter and a smelting reaction slag tank;
  • the smelting reaction device with slag or iron slag can flow out is a plasma furnace, a direct current arc furnace, an alternating current arc furnace, a submerged arc furnace, a blast furnace, a blast furnace, an induction furnace, a cupola furnace, a side blowing molten pool smelting furnace, Bottom-blow pool smelting furnace, top-blow pool smelting furnace, reverberatory furnace, Osmet furnace, Aisa furnace, Waten Kraft melting pool melting furnace, side blowing rotary furnace, bottom blowing rotary furnace, top blowing rotary furnace .
  • the oxide is reduced to metallic copper, metallic nickel, metallic cobalt and FeO, respectively, and the metallic iron content in the slag is ⁇ 3%. It can be regulated by adding one or both of a reducing agent and a carbon-containing iron-containing material, wherein the reducing agent and/or the carbon-containing iron-containing material is used in the amount of copper oxide and nickel oxide in the slag.
  • the theoretical amount of cobalt oxide and iron oxide reduced to metal copper, metallic nickel, metallic cobalt and FeO is 110-140%; the carbon-containing iron-containing material is pre-reduced with steel dust and soot, iron concentrate containing carbon Pellet, iron concentrate carbon-containing metallized pellets, wet zinc smelting kiln slag, coke oven dust and soot.
  • the nickel smelting slag is a nickel smelting slag produced by a "smelting smelting" process, a copper ice nickel blowing process, a blowing slag after a process blowing, and a "copper ice nickel blowing" process.
  • the nickel smelting The slag is in a molten state or a cold state, wherein the molten nickel smelting slag is obtained from a slag outlet of the nickel smelting furnace, or the nickel smelting slag is heated to a molten state;
  • the calcium-based mineral is one or more of lime, limestone, dolomite, calcium carbide slag, red mud or red mud after de-sodium and high calcium; the additive is SiO 2 , MgO, FeO, Fe 2 O 3 , One or more of MnO 2 , Al 2 O 3 , TiO 2 , Fe or Na 2 O.
  • the copper-containing material is blister copper copper fire refining slag, copper tailings, copper slag, zinc smelting slag, zinc smelting soot and dust, lead smelting slag, lead zinc tailings , lead ice copper, arsenic matte, coarse lead fire refining slag, lead smelting soot and soot, lead acid battery, copper smelting soot and dust, copper, copper-containing garbage, copper-containing circuit board, tin smelting slag, tin One or more of tailings; wherein the zinc smelting slag is zinc smelting slag produced by wet zinc smelting and pyrometallurgical smelting, including leaching slag, iron slag slag, copper cadmium slag, goethite slag, Hematite slag, volatile kiln slag, vertical tank zinc slag,
  • the copper smelting slag is produced in the fire smelting process of copper. ⁇ ⁇ ” process; said copper-containing blowing
  • the refining slag is produced in the "copper crucible blowing" process of the pyrometallurgical smelting process of copper; the depleted slag is the copper smelting slag and the copper-containing blowing slag depleted and then discarded, and the flotation tailings are copper-containing smelting slag and The tailings after beneficiation of copper-containing blowing slag.
  • the fuel and the reducing agent are one or more of a solid, a liquid or a gaseous fuel, which are sprayed or fed in a manner of 0 to 1200.
  • the alkaline material is one or more of lime powder, red mud, high calcium red mud, calcium carbide slag, dolomite powder or quicklime powder;
  • the basic iron-containing material is CaO/SiO 2 >1
  • an alkaline sintered ore an alkaline iron concentrate, an alkaline pre-reduced pellet, an alkali metallized pellet, or a steel slag;
  • the acidic material is one or more of silica, gold-silver-silica-containing, fly ash, and coal gangue; and the acidic iron-containing material is CaO/SiO 2 ⁇ 1.
  • iron material acid sinter, acid iron concentrate, acid pre-reduction pellet, acid metallized pellet, copper slag, lead-containing slag, zinc smelting slag, nickel smelting slag, iron alloy slag, blast furnace slagkind.
  • the iron-containing material is ordinary iron concentrate, ordinary iron concentrate direct reduced iron, ordinary iron concentrate sintered ore, ordinary iron concentrate pellet, ordinary iron concentrate metallized ball.
  • ordinary iron concentrate carbon pre-reduction pellets, steel slag, zinc smelting slag, coke smelting soot and dust, steel soot and dust, nickel-containing smelting slag, copper slag, lead smelting slag, copper slag, tin smelting slag
  • the steel dust and dust include blast furnace gas mud, converter dust mud, electric furnace dust, heat (cold Rolling sludge, sintering dust, pellet dust, dust collection in ironworks, blast furnace gas ash, electric furnace dust ash, rolled steel oxide scale;
  • the fluorine-containing material is one or more of fluorite, CaF 2 or fluorine-containing blast furnace slag;
  • the wet zinc slag and dust sludge need to be dehydrated and dried.
  • the solid fuel and the reducing agent are one or more of coal powder, coke powder, coke, fly ash, bituminous coal or anthracite coal, and the shape is granular or powder, and the granular material has a particle size of 5 to 25 mm, and the powder The granular material has a particle size of ⁇ 150 ⁇ m; the liquid fuel and the reducing agent are heavy oil, and the gaseous fuel and reducing agent are gas and/or natural gas.
  • the mixing is uniform or natural mixing or stirring
  • the stirring mode is argon stirring, nitrogen stirring, nitrogen-argon mixed gas stirring, reducing gas stirring, oxidizing gas stirring, electromagnetic One or several of stirring or mechanical agitation.
  • the copper-containing material, the iron-containing material, and the fluorine-containing material are pellets or powdery materials or granulation; wherein, the granular material has a particle size of ⁇ 150 ⁇ m, and the granular material has a particle size of 5 ⁇ . 25mm, the powdery material is sprayed in by spraying, and the granular material is added by spraying or feeding.
  • the loading gas used for the spraying is preheated argon, nitrogen, reducing gas (gas and/or One or more of natural gas) and oxidizing gas, and the preheating temperature is 0 to 1200 °C.
  • the copper-containing material and the iron-containing material are in a hot state or a cold state, wherein the hot material is directly obtained from a metallurgical furnace discharge port or a slag outlet.
  • the copper-rich nickel phase settled at the bottom, the iron-rich phase in the middle portion, and the upper iron-containing silicate mineral phase may be separately treated, Or the middle and upper iron-rich and the iron-containing silicate mineral are combined, the gold component and the silver component migrate and enrich and enter the copper-rich nickel phase, and some of the nickel-cobalt component enters the iron-rich phase, and the zinc-containing component and the The lead component volatilizes and enters the soot recovery as an oxide.
  • the separation and recovery in the step S2 is performed by any one of the following methods 1 to 5:
  • Method 1 When the slag can be used to flow out of the smelting reaction device, after the slag separation after the reaction is completed, the following steps are performed:
  • Method A directly used as a cement raw material after water quenching or air cooling;
  • Method B part or all of the iron-containing silicate mineral phase is returned to the reaction slag as a hot metallurgical flux
  • Method C for pouring glass ceramics or as slag wool
  • Method D slag containing iron silicate mineral phase is oxidized after air cooling or water quenching
  • the method comprises: slag remaining in the smelting reaction device or melting
  • the slag is poured into the heat preservation device, and the preheated oxidizing gas having a temperature of 0 to 1200 ° C is blown into the slag containing iron silicate, and the slag temperature of the silicate is ensured to be >1450 ° C;
  • the weight percentage of iron is ⁇ 1%, and the slag after oxidation is obtained; the slag after oxidation is directly air-cooled or water-quenched, and used as slag cement, cement conditioner, additive in cement production or cement clinker;
  • the preheated fuel and the preheated oxidizing gas are injected, the heat is burned, the heat is supplemented, or the device itself is heated to make the temperature of the iron-containing silicate slag >1450 ° C;
  • Method E The iron-containing silicate mineral phase is used to produce high value-added cement clinker, including the following steps:
  • the iron-containing silicate mineral phase is retained in the smelting reaction device or the slag is poured into the heat preservation device, and molten slag, lime, limestone, iron alloy slag, and molten slag are added to the slag containing the iron silicate mineral phase.
  • molten slag, lime, limestone, iron alloy slag, and molten slag are added to the slag containing the iron silicate mineral phase.
  • fly ash, alkaline iron ore, bauxite, molten blast furnace slag, red mud, red mud after desoda or calcium carbide slag thoroughly mixed to obtain a slag mixture;
  • the slag mixture material is blown into the oxidizing gas with a preheating temperature of 0-1200 ° C, and the slag mixture material temperature is >1450 ° C; when the ferrous oxide weight percentage content is ⁇ 1%, the oxidation is obtained.
  • the oxidized slag is subjected to air cooling or water quenching to obtain a high value-added cement clinker;
  • Method F the slag containing the iron silicate mineral phase as a blast furnace ironmaking raw material or a direct reduction ironmaking raw material: after the iron-containing silicate slag is air-cooled, water-quenched or slowly cooled, used as blast furnace ironmaking or Direct reduction of ironmaking raw materials, after direct reduction, magnetic separation or electric furnace melting, magnetic separation products are metal iron and tailings, electric furnace melting, the product is molten iron and slag;
  • the slag is modified and magnetically separated, including: preheating the oxidizing gas at 0 to 1200 ° C into the slag in the heat preservation device, and ensuring the slag temperature thereof. >1250 ° C;
  • the slag temperature is ⁇ 1250 ° C
  • the preheated fuel and the preheated oxidizing gas are injected, the heat is burned, the heat is supplemented, or the device itself is heated to make the slag temperature >1250 ° C;
  • the oxidized slag is slowly cooled to room temperature, crushed and magnetically selected, and the product is magnetite concentrate and tailings, and tailings are used as building materials;
  • Method G The iron-containing silicate mineral phase is subjected to reduction ironmaking, comprising the following steps:
  • the iron-containing silicate mineral phase is retained in the smelting reaction device or the slag is poured into the heat preservation device, or the iron-containing material is added, and the reducing agent is added at the same time to perform smelting reduction, and the reaction slag is monitored in real time, and the slag is controlled.
  • the method of controlling the temperature of the reaction slag is:
  • a metallurgical flux is added to the reaction slag, so that the temperature of the reaction slag reaches 1350 to 1670 ° C, wherein
  • the metallurgical flux is a mineral containing CaO or SiO 2 , specifically one or more of quartz sand, gold-silver quartz sand, red mud, high-calcium red mud after desodiumification, calcium carbide slag, dolomite or limestone;
  • the method of controlling the alkalinity of the reaction slag is:
  • Method I Pour the mixed slag after reduction into a heat preservation slag tank, and slowly cool to room temperature to obtain slow cooling slag; wherein, the metal iron settles to the bottom of the reaction device to form iron slag, and the remaining slow slag contains metal iron Layer, broken to a particle size of 20 ⁇ 400 ⁇ m, grinding, magnetic separation to separate the remaining metal iron and tailings;
  • Method II the mixed slag after reduction, cooling and sedimentation, separation of slag-gold, obtaining molten iron and reduced slag; and the slag after reduction is melted according to one or several methods of methods A to E Slag treatment; the molten iron is sent to a converter or an electric furnace for steelmaking;
  • part of the gold-containing component and the silver-containing component enter the copper-rich nickel phase, and the nickel and cobalt components respectively enter the copper-rich nickel phase and the iron-rich phase;
  • the iron-rich phase layer is obtained by water quenching or air cooling or pouring into a heat preservation device, or after being manually sorted and re-selected, as a blast furnace smelting nickel-iron raw material or directly reducing ferronickel raw material or Smelting reduction of nickel-iron raw materials or flotation of copper raw materials; during flotation, the flotation products are copper concentrate, nickel concentrate, nickel-containing alloy and iron concentrate, iron concentrate as blast furnace ironmaking raw material or direct reduction Iron raw material or smelting reduction ironmaking raw material; wherein the direct reduction process uses a rotary hearth furnace, a tunnel kiln, a vehicle bottom road, a shaft furnace, a rotary kiln or an induction furnace as a reduction device, using gas-based or coal-based reduction technology, gas The base is reduced by natural gas and/or gas, and the coal-based reduction adopts one or more of anthracite, bituminous coal, lignite, coking coal, coke powder or coke, and the controlled reduction
  • red mud contains potassium, sodium, dust, and steel soot containing lead, zinc, antimony, and indium, when these materials are added, some indium, antimony, potassium, and sodium components are added. Volatile, entering the soot as an oxide.
  • Method 2 When the smelting reaction device through which the slag can flow out is used, the obtained iron-rich phase and the iron-containing silicate mineral phase treatment method are treated by one or more of the methods A to G described in the first method, Pour the copper-rich nickel phase after the slow cooling of the heat preservation device, and send it to a converter or a rotary furnace or as a raw material for copper, cobalt and nickel separation.
  • Method 3 using the slag rotatable converter and the reaction slag tank, obtaining the molten iron-containing silicate mineral phase, and the treatment method is treated by one or more of the methods A to G described in the first method; Or containing the iron-rich phase by the step S2-1-05 in the first method; the molten state or the copper-rich nickel phase after being poured into the heat preservation device to be cooled, sent to a converter or a rotary furnace or as copper cobalt nickel Separated raw materials.
  • Method 4 When the slag rotatable converter and the reaction slag tank are used, the molten iron-containing silicate mineral phase and the iron-rich phase obtained in the molten state are obtained, and the treatment method is one of the methods A to G described in the first method or The treatment is carried out in several ways; the molten copper-rich nickel phase is sent to a converter or a blowing furnace for copper smelting, or after slow cooling, the metal iron is separated by magnetic separation and then sent to a converter or a converting furnace or as a copper-cobalt-nickel separation. raw material.
  • Method 5 When using a heat preservation device, or using a smelting reaction device through which slag can flow out, when pouring the slag into the heat preservation device, perform the following steps:
  • the slag is cooled to room temperature to obtain slow cooling slag; the nickel-rich copper phase is settled to the bottom of the reaction device to form copper-rich nickel ruthenium; the iron-containing silicate mineral phase floats; the middle is slow cooling
  • the slag is an iron-rich phase, and at the same time, a zinc-containing component and a lead-containing component are formed; wherein the gold and silver components migrate to the copper-rich nickel phase, and the nickel and cobalt components respectively migrate to the copper-rich nickel phase and the iron-rich phase;
  • the oxidizing gas is one of preheated air, oxygen, oxygen-enriched air, nitrogen-air, argon-air, oxygen-nitrogen, oxygen-argon.
  • the mixing is uniformly mixed by natural mixing and stirring, and the stirring and mixing is performed by argon stirring, nitrogen stirring, nitrogen-argon gas mixture stirring, and reducing property.
  • gas agitation oxidative gas agitation, electromagnetic agitation or mechanical agitation;
  • the sedimentation is natural sedimentation or spin sedimentation or centrifugal sedimentation;
  • the cooling mode when cooling sedimentation is natural cooling or rotary cooling or centrifugal cooling, and the gravity sorting method is a shaker. Sorting, chute sorting or a combination of the two.
  • the method for recovering valuable components from the nickel-containing smelting slag of the present invention can treat the hot slag, fully utilize the physical heat resources of the molten nickel smelting slag and the hot metallurgical flux, and can also treat the cold slag.
  • the metallurgical process of nickel-containing slag is realized; and the current accumulation of slag, environmental pollution, and heavy metal element pollution are solved;
  • the slag metallurgical reaction in the slag is added with calcium minerals to disintegrate the olivine, and the iron oxide is fully released to form free iron oxide, which realizes the growth and settlement of the iron-rich phase, and the slag
  • the iron-containing component aggregates, grows and settles, and the iron-rich phase includes a plurality of metal iron, FeO phase, and fayalite phase, and is used as a raw material for blast furnace ironmaking or direct reduction or smelting reduction ironmaking;
  • the calcium-based minerals effectively modify the viscosity, which is a decrease in viscosity and contribute to the sedimentation of the copper-containing component.
  • the copper component, nickel component, cobalt component, gold and silver component in the mixed slag migrate, enrich and concentrate in the copper-rich nickel phase, and achieve growth and sedimentation, wherein the copper-rich nickel phase includes copper , white ice copper, copper ice nickel, copper-rich nickel-cobalt phase, a plurality of iron-containing components, or part of the copper-nickel-cobalt component enters the iron-rich phase;
  • the slag can be tempered and can be used as cement raw materials or building materials or instead of crushed stone as aggregate and road material;
  • the method of the invention adds an additive, one is used for reducing the viscosity, the other is for lowering the melting point, and the copper-rich nickel phase is precipitated at a certain temperature (1100 to 1450 ° C), and the iron is obtained after the sedimentation is separated.
  • the phase is a low copper iron-rich phase and an iron-containing silicate phase, wherein the iron-rich phase and the iron-containing silicate phase have a copper content of less than 0.1%, and the iron concentrate can be obtained by direct reduction or smelting reduction of iron.
  • the method of the present invention can be carried out continuously or intermittently to meet the needs of industrial production.
  • the raw material of the present invention may be cold slag, cold slag treatment, not only can realize the copper component in the slag,
  • the comprehensive utilization of the nickel component, cobalt component, gold component, silver component, strontium component, sodium component, potassium component, iron component, zinc component, and lead component valuable component can solve the current A large amount of slag is accumulated and environmental pollution problems.
  • the raw material of the present invention may be liquid molten nickel smelting slag flowing out of the slag outlet, which contains abundant thermal energy resources, has the characteristics of high temperature and high heat, fully utilizes the slag physical heat resource, and efficiently saves energy;
  • the molten copper slag contains a large amount of hot metallurgical flux, and is a slag system with excellent physical and chemical properties, realizing slag metallurgy.
  • the invention adjusts the physicochemical properties of the slag, blows the gas, controls the oxygen potential, and causes the copper component, the nickel component, the cobalt component, the gold and silver component in the slag to migrate and enrich to the copper-rich nickel phase. To achieve aggregation, growth and settlement.
  • adding cold material and molten nickel smelting slag avoids excessive slag temperature and improves the life of the heat preservation device; adding cold material and nickel smelting slag improves the processing amount of the raw material, and can not only process the liquid melting Slag, and can handle a small amount of cold materials, the raw material adaptability; the addition of cold materials to achieve the chemical heat released by the reaction and the efficient use of slag physical heat.
  • the invention adjusts the physicochemical properties of the slag, controls the oxygen potential, and simultaneously adds the calcium-based minerals to release the iron oxides in the olivine phase, enriched in the iron-rich phase, and achieves aggregation, growth and sedimentation, and iron-rich
  • the phase includes a plurality of metal iron, FeO phase, and fayalite phase, and is used as a raw material for blast furnace ironmaking or direct reduction or smelting reduction of iron; zinc component, lead component, indium component, strontium component, sodium in slag
  • the components and potassium components are volatilized and enter the soot for recycling.
  • the addition of the additive, the copper component, the nickel component, the cobalt component, the gold and silver component in the slag are enriched in the copper-rich nickel phase, and the aggregation, growth and sedimentation are realized.
  • the copper-rich nickel phase comprises copper, white copper, copper ice nickel, copper nickel cobalt phase, a plurality of iron-containing components, or a part of the copper-nickel-cobalt component enters the iron-rich phase, and the iron component is enriched in the slag.
  • the slag-containing insulation device is placed on the rotating platform to accelerate the accumulation, growth and settlement of the copper-rich nickel phase and the iron-rich phase; Accelerate the growth and settlement of the copper-rich nickel phase and the iron-rich phase.
  • the method of the invention adopts the methods of manual sorting, magnetic separation, re-election or slag-gold separation to respectively distribute the low-iron silicate mineral phase, the iron-rich phase and the copper-rich nickel phase distributed in the upper part, the middle part and the bottom part. Separation is carried out to achieve efficient recovery of copper component, nickel component, cobalt component, gold and silver component, indium component, antimony component and iron component in the slag, and the obtained iron-containing silicate phase and iron-rich phase are obtained.
  • the slag of the component contains ⁇ 0.1wt% copper.
  • the amount of slag to be sorted is small, the mineral grindability is increased, and the cost of grinding, magnetic separation and re-election is increased. Low; the subsequent separation process uses physical beneficiation (magnetic separation or re-election), the separated medium is water, and there is no environmental pollution during the separation process.
  • the slag treatment process has short process, simple operation, high recovery rate, and high efficiency. Clean and environmentally friendly; tailings are used as cement raw materials, building materials, instead of crushed stone as aggregates and road materials.
  • the present invention precipitates by the copper-rich nickel phase, and the iron-containing silicate phase and the iron-rich component have a copper content of less than 0.1%, and are used as a raw material for blast furnace ironmaking or direct reduction or smelting reduction to obtain metallic iron and molten iron.
  • the invention fully utilizes the slag physical heat resource and the hot metallurgical flux, and can also treat the cold slag, realizes the slag metallurgy, the copper component, the nickel component, the cobalt component, the gold and silver group in the slag It is enriched in the copper-rich nickel phase and achieves aggregation, growth and sedimentation.
  • the iron component is enriched in the iron-rich phase to achieve aggregation, growth and sedimentation. Manual sorting, magnetic separation, re-election and slag-gold are used.
  • the combination method of sedimentation combines the copper-rich nickel phase and the iron-rich phase settled in different parts to realize efficient recovery of copper component, nickel, cobalt, gold and silver components and iron components in the slag;
  • the method has the advantages of short process flow, high metal recovery rate, low production cost, strong adaptability of raw materials, large processing capacity, environmental friendliness and high economic benefit, and can solve the problem of efficient recycling of metallurgical resources and heat energy.
  • a method for recovering valuable components from a smelting slag containing nickel comprising the following steps:
  • Step 1 slag mixing:
  • the smelting slag containing nickel is added to the smelting reaction device through which the heat preservation device or the slag can flow out, and the calcium-based mineral and the additive are added to form a mixed slag;
  • the mixed slag is heated to a molten state to form a reaction slag; the mixture is uniformly mixed, and the nickel-containing reaction slag is monitored in real time, and the slag after completion of the reaction is obtained by controlling two parameters (a) and (b) as follows. Or pouring the slag after the completion of the reaction into the heat preservation device;
  • the control method is:
  • the method for controlling the temperature of the reaction slag in the set temperature range is:
  • the heating function of the reaction device itself, or the reaction slag is added with fuel and preheated oxidizing air, so that the temperature of the reaction slag reaches a set temperature range;
  • reaction slag When the temperature of the reaction slag is below the upper limit of the set temperature range, one of nickel smelting slag, blast furnace slag, iron alloy slag, steel slag, copper-containing material, metallurgical flux, iron-containing material or fluorine-containing material is added to the reaction slag. Or several, so that the temperature of the mixed slag reaches a set temperature range;
  • the ratio of alkalinity CaO/SiO 2 in the reaction slag is ⁇ 0.15, one or more of an alkaline material or an alkaline iron-containing material is added to the reaction slag;
  • reaction slag When the ratio of alkalinity CaO/SiO 2 in the reaction slag is >1.5, one or more of an acidic material or an acidic iron-containing material is added to the reaction slag;
  • the slag After 5 to 50 minutes of heat preservation, the slag after the completion of the reaction, sedimentation, and slag-gold separation, obtain the copper-rich nickel in the bottom molten state, the iron-rich phase in the middle and lower molten state, and the iron-containing silicate mineral phase in the middle and upper molten state. Simultaneously forming a zinc-containing component and a lead-containing component, wherein the indium, antimony, nickel, cobalt, gold, and silver components migrate to the nickel-rich copper phase, and part of the nickel-cobalt component migrates to the iron-rich phase;
  • Each phase is recovered using one of the following methods:
  • the iron-rich phase is obtained by water quenching or air cooling or pouring into a heat preservation device, or by manual sorting and re-election, as a raw material for blast furnace ironmaking or direct reduction of nickel-iron raw materials or smelting reduction of nickel-iron raw materials or float Selecting copper and nickel raw materials; flotation products are copper concentrate, nickel concentrate, nickel-containing alloy and iron concentrate, iron concentrate as blast furnace ironmaking raw material or direct reduction ironmaking raw material or smelting reduction ironmaking raw material;
  • the reduction process uses a rotary hearth furnace, a tunnel kiln, a vehicle bottom road, a shaft furnace, a rotary kiln or an induction furnace as a reduction device, and the gas-based or coal-based reduction technology is used to reduce the gas base to natural gas and/or gas, and the coal base is reduced to anthracite.
  • One or more of bituminous coal, lignite, coking coal, coke breeze or coke, the reduction temperature is 900-1400 ° C,
  • the iron-containing silicate mineral phase in the step (1) is subjected to slag treatment, and one of the methods A to G is adopted:
  • Iron-containing silicate mineral phase as cement raw material Iron-containing silicate mineral phase as cement raw material
  • the iron-containing silicate mineral phase is directly quenched or air-cooled as a cement raw material or further processed into a high value-added cement raw material.
  • Method B Part or all of the iron-containing silicate mineral phase is returned to the copper-containing reaction slag:
  • Some or all of the iron-containing silicate mineral phase is returned to the copper-containing reaction slag, and as a hot metallurgical flux, the copper-containing reaction slag component is adjusted to control the copper-containing reaction slag temperature.
  • Method C pouring a glass-ceramic with a ferrite-containing mineral phase or as a slag wool.
  • Method D Air-cooling or water quenching after oxidation of iron-containing silicate slag:
  • the preheated fuel and the preheated oxidizing gas are injected, the heat is burned, the heat is supplemented, or the device is heated by itself, so that the silicate slag temperature is >1450 ° C;
  • the slag after oxidation is directly air-cooled or water-quenched, and used as slag cement, cement conditioner, additive in cement production or cement clinker.
  • Method E Treatment of high value-added cement clinker by treatment with iron silicate slag:
  • Method F the iron-containing silicate mineral phase slag is used as a blast furnace ironmaking raw material or a direct reduction ironmaking raw material: the slag containing the iron silicate mineral phase is air-cooled, water quenched or slowly cooled, and used as a blast furnace Iron or direct reduction of ironmaking raw materials, after direct reduction, magnetic separation or electric furnace melting, magnetic separation products are metal iron and tailings, electric furnace melting, the product is molten iron and slag;
  • the separation is performed by the following method: magnetic separation after slag modification: slag flowing into the heat preservation device, blowing 0-1200 ° C Preheating the oxidizing gas, and ensuring that the slag temperature is >1250 ° C, completing the transformation of magnetite in the slag; slowly cooling the oxidized slag to room temperature, crushing, magnetic separation, and the product is magnetite Mines and tailings, tailings as building materials.
  • the control method is:
  • the method for controlling the temperature of the reaction slag in the set temperature range is:
  • the temperature of the reaction slag ⁇ lower limit of the set temperature range is 1350 ° C
  • the temperature of the reaction slag reaches the set temperature range by the heating function of the reaction device itself or by adding the fuel and the preheated oxidizing gas to the slag.
  • the temperature range and the alkalinity control method are the same as the method G step (1);
  • Method I Perform the following steps:
  • Method II Perform the following steps:
  • the specific method is: using one or more of the methods A to E in the separation and recovery method 1 of the step 2, the slag treatment;
  • the gas produced by the reduction is secondarily burned on the surface of the slag to provide heat, and the gas flowing out of the furnace can be used as a heat source for the drying charge and the heat preservation device.
  • the molten iron-containing silicate mineral phase is subjected to slag treatment by a slag treatment by one or more of the methods A to G in the separation and recovery method 1 of the step 2;
  • the step of containing the iron-rich phase is carried out by the step (4) in the first method;
  • the molten iron-containing silicate mineral phase and the iron-rich phase are subjected to slag treatment in a specific manner: one or more of the methods A to G in the separation and recovery method 1 of the step 2 are used;
  • Method 5 When using a heat preservation device, or using a smelting reaction device through which slag can flow out, when the slag is poured into the heat preservation device, the slag after the reaction is completed is as follows:
  • the direct reduction process uses a rotary hearth furnace, a tunnel kiln, a vehicle bottom road, a shaft furnace, a rotary kiln or an induction furnace as a reduction device, using gas-based or coal-based reduction technology, and gas-based reduction using natural gas and/or gas, coal-based
  • the reduction adopts one or more of anthracite, bituminous coal, lignite, coking coal, coke powder or coke, and the controlled reduction temperature is 900-1400 ° C, and the control alkalinity CaO/SiO 2 ratio is 0.8-1.5;
  • the nickel smelting slag is in a molten state or a cold state, wherein: the molten nickel smelting slag is obtained from a slag opening of a nickel smelting furnace, or the nickel smelting slag is heated To the molten state; the nickel smelting slag is the nickel smelting slag produced by the "smelting smelting" process, the copper smelting nickel blowing process, the blowing slag after the process blowing, and the blowing slag after the "copper ice nickel blowing" process blowing Depleted slag produced by depletion, nickel smelting slag produced by top-blown smelting, and one or more of sedimentation slag produced by sedimentation of nickel smelting slag produced by top-blown smelting;
  • the copper materials in steps 1 and 2 are crude copper fire refining slag, copper slag, copper tailings, zinc smelting slag, zinc smelting soot and dust, lead smelting slag, lead zinc tailings, lead copper, Arsenic matte, coarse lead fire refining slag, lead smelting soot and dust, lead acid battery, copper smelting soot and dust, copper, copper-containing garbage, copper-containing circuit board, tin smelting slag, tin tailings
  • zinc smelting slag is zinc smelting slag produced by wet zinc smelting and pyrometallurgical smelting, including leaching slag, iron slag, copper cadmium slag, goethite slag, hematite slag, volatile kiln residue , vertical tank zinc slag, blast furnace slag, smelting furnace slag, vortex furnace s
  • the copper slag is one or more of copper-containing smelting slag, copper-containing blowing slag, depleted slag, flotation tailings, and wet copper slag, and copper smelting slag
  • the smelting reaction device through which the slag can flow is a rotatable smelting reaction device or a smelting reaction device with a slag port or an iron port;
  • the heat preservation device is a pourable smelting reaction slag irrigation and heat preservation pit
  • the rotatable smelting reaction device is a converter and a smelting reaction slag tank;
  • the smelting reaction device with the slag port or the iron slag flowing out is a plasma furnace, a direct current arc furnace, an alternating current arc furnace, a submerged arc furnace, a blast furnace, a blast furnace, an induction furnace, a cupola, and a side blowing molten pool melting furnace.
  • the calcium-based mineral is specifically one or more of lime, limestone, dolomite, calcium carbide slag, red mud or post-sodium high calcium red mud;
  • the additive is one or more of SiO 2 , MgO, FeO, Fe 2 O 3 , MnO 2 , Al 2 O 3 , TiO 2 , Fe or Na 2 O.
  • the two parameters (a) and (b) are ensured at the same time, and the copper oxide, nickel oxide, cobalt oxide and iron oxide in the slag are respectively reduced to metal copper and metal.
  • Nickel, metallic cobalt and FeO, the metal iron content in the slag is ⁇ 3%. It can be regulated by adding one or both of a reducing agent and a carbon-containing iron-containing material, wherein the reducing agent and/or the carbon-containing iron-containing material is used in the amount of copper oxide and nickel oxide in the slag.
  • the theoretical amount of cobalt oxide and iron oxide reduced to metal copper, metallic nickel, metallic cobalt and FeO is 110-140%; the carbon-containing iron-containing material is pre-reduced with steel dust and soot, iron concentrate containing carbon Pellet, iron concentrate carbon-containing metallized pellets, wet zinc smelting kiln slag, coke oven dust and soot.
  • the fuel and the reducing agent are one or more of a solid, liquid or gaseous fuel, which is sprayed or charged, and the loaded gas is preheated and oxidized.
  • One or more of gas, nitrogen or argon, preheating temperature is 0-1200 ° C;
  • solid fuel and reducing agent is one or more of coal powder, coke powder, coke, fly ash, bituminous coal or anthracite Species, the shape is granular or powder, the granular material has a particle size of 5 to 25 mm, the powdery material has a particle size of ⁇ 150 ⁇ m, the liquid fuel and the reducing agent are heavy oil, and the gaseous fuel and the reducing agent are gas and/or natural gas;
  • the metallurgical flux in the steps 1 and 2 is a mineral containing CaO or SiO 2 , specifically quartz sand, gold-silver-sand quartz sand, red mud, high-calcium red mud after desoda, calcium carbide slag, dolomite or limestone.
  • quartz sand specifically quartz sand, gold-silver-sand quartz sand, red mud, high-calcium red mud after desoda, calcium carbide slag, dolomite or limestone.
  • quartz sand specifically quartz sand, gold-silver-sand quartz sand, red mud, high-calcium red mud after desoda, calcium carbide slag, dolomite or limestone.
  • the iron-containing materials in 1 and 2 are ordinary iron concentrate, ordinary iron concentrate direct reduced iron, ordinary iron concentrate sintered ore, ordinary iron concentrate pellet, ordinary iron concentrate metallized pellet, ordinary iron Concentrate carbon pre-reduction pellets, steel slag, zinc smelting slag, coke smelting soot and dust, steel soot and dust, nickel-containing smelting slag, copper slag, lead smelting slag, zinc smelting slag, tin smelting slag, red mud
  • the steel dust and dust include blast furnace gas mud, converter dust mud, electric furnace dust, hot (cold) smelting Mud, sintered dust, pellet dust, dust collection in ironworks, blast furnace gas ash, electric furnace dust ash, steel oxide scale.
  • the copper-containing material and the iron-containing material are in a hot or cold state, wherein the hot material is directly obtained from a metallurgical furnace discharge port or a slag outlet.
  • the wet zinc slag and dust are subjected to dehydration and drying.
  • zinc smelting slag and soot, lead smelting slag and soot contain indium and antimony, lead, silver, zinc and antimony; red mud contains sodium and potassium, and steel soot and dust contain indium, antimony and silver.
  • Sodium and potassium the above materials all have iron, lead smelting slag and zinc smelting slag contain copper, copper soot and dust contain indium and antimony, so in the method of the invention, indium, antimony, sodium, potassium, zinc, lead will The fumes are entered in the form of oxides for recycling.
  • the fluorine-containing materials in 1 and 2 are one or more of fluorite, CaF 2 or fluorine-containing blast furnace slag;
  • the copper-containing material, the iron-containing material and the fluorine-containing material are pellets or powdery materials or granulation; wherein, the granular material has a particle size of ⁇ 150 ⁇ m, and the granular material has a particle size of 5 to 25 mm.
  • the powdery material is sprayed by spraying, and the granular material is added by spraying or feeding, and the loading gas is preheated argon gas, nitrogen gas, reducing gas (gas and/or natural gas), oxidizing gas.
  • One or more, preheating temperature is 0 to 1200 ° C;
  • the injection method is adopted, and the refractory spray gun is specifically inserted into the slag or placed.
  • the refractory spray gun is specifically inserted into the slag or placed.
  • the copper component, the nickel component, the cobalt component, the gold and silver component in the slag are enriched in the copper-rich nickel phase, and the aggregation, growth and sedimentation are achieved.
  • the iron component is enriched in the iron-rich phase to achieve aggregation, growth and sedimentation, and the zinc component, the lead component, the indium component and the strontium component in the slag are separately collected into the soot.
  • the temperature of the mixed slag is > the upper limit of the set temperature
  • one or more of nickel smelting slag, copper-containing material, metallurgical flux, iron-containing material or fluorine-containing material are added, in order to avoid excessive temperature and protect the refractory material.
  • Another effect of adding fluorine-containing materials is to lower the viscosity, accelerate the aggregation, growth and sedimentation of the copper-rich nickel phase and the iron-rich phase in the slag.
  • the alkaline material is one or more of lime powder, red mud, red mud after desoda, calcium carbide slag, dolomite powder or quicklime powder;
  • the alkaline iron-containing material is an iron-containing material having CaO/SiO 2 >1, and the basic iron-containing material is an alkaline sintered ore, an alkaline iron concentrate, an alkaline pre-reduction pellet, and an alkali metallized sphere.
  • a pellet, steel slag or blast furnace slag is one or more of lime powder, red mud, red mud after desoda, calcium carbide slag, dolomite powder or quicklime powder
  • the alkaline iron-containing material is an iron-containing material having CaO/SiO 2 >1
  • the basic iron-containing material is an alkaline sintered ore, an alkaline iron concentrate, an alkaline pre-reduction pellet, and an alkali metallized sphere.
  • the acidic material is one or more of silica, fly ash and coal gangue;
  • the acidic iron-containing material is an iron-containing material with CaO/SiO 2 ⁇ 1, acidic
  • the iron material is acid sinter, acid iron concentrate, acid pre-reduction pellet, acid metallized pellet, copper slag, lead smelting slag, zinc smelting slag, nickel smelting slag, tin smelting slag, iron alloy slag, blast furnace slag One or several;
  • step 1 while ensuring the two parameters (a) and (b), the slag is thoroughly mixed, and the mixing mode is natural mixing or stirring mixing, and the stirring mode is one of the following modes: argon stirring, nitrogen gas.
  • argon stirring nitrogen gas.
  • One or more of stirring, nitrogen-argon mixed gas agitation, reducing gas agitation, oxidizing gas, electromagnetic stirring, mechanical agitation, gas preheating temperature is 0 to 1200 ° C;
  • the copper-rich nickel phase and the iron-rich phase in the slag are aggregated, grown and settled, which is favorable for the silicate to float.
  • the copper-rich nickel phase includes copper, white ice copper, copper ice nickel, and copper rich.
  • a nickel-cobalt phase, a plurality of iron-containing components, or a part of the copper-nickel-cobalt component enters the iron-rich phase, and the iron-rich phase includes a plurality of metal iron, FeO phase, and fayalite phase;
  • the oxidizing gas is one of preheated air, oxygen or oxygen-enriched air
  • the blowing method is to insert the slag into the molten slag or to be placed on the upper part or the side or the bottom of the reaction slag.
  • the preheating temperature is 0 to 1200 ° C;
  • the copper-containing material and the iron-containing material are in a hot state or a cold state, and the hot material is a hot material directly produced from a metallurgical furnace, and the temperature of the hot material is 200 to 1750. °C;
  • step 2 when cooling and sedimentation is required, the cooling mode is natural cooling or rotary cooling or centrifugal cooling, and the sedimentation mode is natural sedimentation or rotary sedimentation or centrifugal sedimentation;
  • the specific operation of the rotation and the centrifugation is: the device containing the slag after the reaction is completed is placed on the rotating platform, and is rotated according to a certain speed, and the rotation speed is determined according to the quality of the slag and the height or depth of the heat preservation device, and the rotation time Depending on the quality of the slag and the solidification of the slag; the device containing the slag after completion of the reaction is placed on a rotating platform for the purpose of accelerating the accumulation of copper-rich nickel phase, iron-rich phase, growth and sedimentation, which is beneficial to The silicate floats up, shortens the settling time, improves the settlement effect, and improves the production efficiency;
  • step 2 in the slag cooling process after the completion of the reaction, most of the copper-rich nickel phase and the iron-rich phase settle in the middle and lower portions due to the difference in density and mineral size;
  • the copper component, the nickel component, the cobalt component, the gold and silver component in the slag after the reaction is completed
  • the copper component, the nickel component, the cobalt component, the gold and silver component in the slag continues to be enriched in the iron-rich phase, and achieves growth and sedimentation.
  • the gravity sorting method when separating, may be a shaker sorting, a chute sorting or a combination of the two.
  • the iron-rich phase and the iron-containing silicate phase obtained in the slag contain copper ⁇ 0.1%, and the final iron recovery rate is ⁇ 91%, and the zinc recovery rate is ⁇ 92%.
  • the recovery rate is ⁇ 92%
  • the nickel enrichment rate is ⁇ 91%
  • the cobalt enrichment rate is ⁇ 91%
  • the gold enrichment rate is ⁇ 90%
  • the silver enrichment rate is ⁇ 90%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing:
  • the molten nickel smelting slag obtained from the slag smelting process of the smelting process is added to a DC arc furnace, and lime, and SiO 2 , MgO, and Al 2 O 3 are added to form a mixed slag; the mixed slag is heated to melt State, form reaction slag, and make the reaction slag achieve natural mixing; monitor the reaction slag in real time, and ensure the slag after completion of the reaction by regulating and simultaneously satisfying two parameters (a) and (b); corresponding to (a):
  • the temperature of the reaction slag is 1660 ° C, and the refractory spray gun is inserted into the reaction slag.
  • the nitrogen gas is used as the loading gas, and the copper slag, the copper-containing soot, the copper-containing soot and the copper-containing garbage and the copper containing powder having a normal particle size ⁇ 150 ⁇ m are sprayed.
  • Step 2 separation and recovery method 1:
  • the slag after the completion of the reaction is self-precipitated, and the slag-gold is separated to obtain a molten copper-rich nickel layer, an iron-rich layer and an iron-containing silicate mineral phase, and at the same time, a zinc component and a lead component are formed into the soot.
  • the following steps are as follows: (1) the molten iron-containing silicate mineral phase is subjected to the slag treatment outside the furnace, and the method F is used. After the silicate slag is air-cooled, it is used as a direct reduction ironmaking raw material, and in the direct reduction process, the slewing is adopted. As a reduction equipment, the kiln uses gas-based reduction technology.
  • the gas-based reducing agent is natural gas and coal gas.
  • the reduction temperature is 950 ° C
  • the alkalinity CaO / SiO 2 ratio is 0.8
  • the electric furnace melting temperature is 1550 ° C after reduction.
  • the product is metal.
  • the slag contains copper ⁇ 0.1%
  • the zinc recovery rate is 92%
  • the lead recovery rate is 93%
  • the indium recovery rate is 92%
  • cesium recovery rate is 94%
  • sodium recovery rate is 95%
  • potassium recovery rate is 96%, iron 96% yield, 93% enrichment of nickel, cobalt enrichment was 95%, 93% enrichment of gold, silver was 9
  • the copper content of the slag refers to the slag phase after the phase separation of the copper-rich nickel phase, specifically the copper content in the iron-rich phase and the silicate mineral phase, and the enrichment ratio of nickel and cobalt. It refers to the content of nickel and cobalt in the copper-rich nickel phase as a percentage of the total amount of nickel and cobalt in the raw material.
  • the enrichment ratio of gold and silver means that the content of gold and silver in the copper-rich nickel phase accounts for gold and silver in the raw material. The percentage of the total.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing:
  • the molten nickel smelting slag obtained from the slag outlet of the "copper ice nickel blowing" process is added to the pourable smelting reaction slag, and limestone, dolomite, red mud and FeO and Fe 2 O 3 are added to form a mixed melting.
  • Slag using an oxygen-enriched air with a preheating temperature of 600 ° C and a particle size of 20 mm anthracite and coke particles, and blowing natural gas, heating the mixed slag to a molten state to form a copper-containing reaction slag; monitoring the reaction slag in real time, through regulation At the same time, the two parameters (a) and (b) are guaranteed, and the slag after the completion of the reaction is obtained;
  • copper-containing reaction slag temperature is 1660 ° C, using a refractory spray gun to insert into the reaction slag, using argon as carrier gas, spraying normal temperature powder particle size ⁇ 150 ⁇ m copper slag, nickel smelting slag, copper-containing soot, miscellaneous Copper and steel sintered dust, sintered pellet dust, iron plant dust, ordinary iron concentrate direct reduced iron, the temperature is reduced to 1480 ° C;
  • copper-containing reaction slag basicity CaO / SiO 2 ratio of 2.4 Adding an acid iron concentrate, an acidic pre-reduction pellet, a lead-containing smelting slag, and a lead-containing smelting furnace slag to the reaction slag to reduce the alkalinity ratio of the copper-containing reaction slag to 1.2; the metal iron content in the slag is 3%;
  • Step 2 separation and recovery method 2:
  • the temperature of the reaction slag is 1460 ° C, in the temperature range;
  • the zinc recovery rate is 95%, the lead recovery rate is 93%, the indium recovery rate is 96%, the hydrazine recovery rate is 96%, the sodium recovery rate is 97%, the potassium recovery rate is 98%, and the slag contains copper ⁇ 0.1%, iron.
  • the recovery rate was 97%; the nickel enrichment rate was 92%, the cobalt enrichment rate was 96%, the gold enrichment rate was 90%, and the silver enrichment rate was 90%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: the molten nickel smelting slag obtained by the "copper ice nickel blowing" process blowing slag through the slag outlet of the depleting furnace is added to the reverberatory furnace, and limestone and high-calcium red mud after de-sodium are added to form a mixture.
  • Slag with an oxygen preheating temperature of 900 ° C, spray anthracite, coke and pulverized coal with a particle size of 20 mm, and heat the mixed slag to the melt Melting state, forming copper-containing reaction slag, mechanically stirring and mixing; monitoring reaction slag in real time, and simultaneously controlling two parameters (a) and (b) to obtain slag after completion of reaction;
  • the temperature of the copper-containing reaction slag is 1685 ° C, adding acid metallized pellets and copper slag to the reaction slag, and simultaneously adding copper-containing soot, lead-containing slag, ordinary iron concentrate pellets, and rolling steel Iron oxide scales and ordinary iron concentrates contain carbon pre-reduction pellets to lower the temperature to 1420 ° C;
  • alkalinity CaO / SiO 2 ratio of copper-containing reaction slag 2.3, adding quartz sand to the reaction slag a mixture of red mud and zinc smelting slag, the alkalinity ratio of the copper-containing reaction slag is reduced to 1.2; the metal iron content in the slag is 1.8%;
  • Step 2 separation and recovery method 2:
  • the zinc-containing component, the indium component, the antimony component and the lead-containing component volatilize and enter the soot recovery.
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 96%, and the zinc recovery rate is 93%.
  • the recovery rate is 92%, the indium recovery rate is 93%, the ruthenium recovery rate is 94%; the nickel enrichment rate is 93%, the cobalt enrichment rate is 97%, and the gold enrichment rate is 91%.
  • the enrichment rate was 92%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: adding molten smelting slag obtained from the slag outlet of the top blowing molten pool to a plasma furnace, simultaneously adding dolomite, MgO, Al 2 O 3 , and Fe to form mixed slag; mixing slag Heating to a molten state, forming a copper-containing reaction slag, and electromagnetically stirring the reaction slag to achieve mixing; monitoring the reaction slag in real time, and simultaneously controlling the two parameters (a) and (b) to obtain a melting after completion of the reaction Slag
  • Step 2 separation and recovery method 2:
  • the slag contains copper ⁇ 0.1%, the recovery of iron is 95%, and the recovery of zinc
  • the rate is 92%, the lead recovery rate is 92%, the indium recovery rate is 93%, the ruthenium recovery rate is 94%; the nickel enrichment rate is 93%, the cobalt enrichment rate is 94%, and the gold enrichment rate At 92%, the silver enrichment rate is 94%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: the molten nickel smelting slag obtained by melting the smelting slag from the top blowing molten pool through the sedimentation electric furnace slag opening is added to the thermal insulation slag tank, and at the same time adding limestone and Fe to form mixed slag; the preheating temperature is 600 ° C The oxygen-enriched air is sprayed with bituminous coal with a particle size of ⁇ 150 ⁇ m, and the mixed slag is heated to a molten state to form a copper-containing reaction slag, and the reaction slag is mixed; the reaction slag is monitored in real time, and the (a) and (b) two parameters to obtain the slag after completion of the reaction;
  • the temperature of the copper-containing reaction slag is 1430 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 1.5, both within the required range;
  • the metal iron content in the slag is 2.2 %;
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 92%, the zinc recovery rate is 94%, and the lead recovery rate is 95. %; nickel enrichment rate is 95%, cobalt enrichment rate is 93%, gold enrichment rate is 91%, and silver enrichment rate is 92%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: molten nickel smelting slag obtained from the slag smelting process of the smelting process, and molten smelting slag obtained from the slag outlet of the "copper ice nickel blowing" process blowing furnace are added to the alternating current arc furnace At the same time, adding lime, MgO, Al 2 O 3 , Fe 2 O 3 to form a mixed slag; heating the mixed slag to a molten state to form a copper-containing reaction slag, and spraying argon gas having a preheating temperature of 800 ° C, And the reaction slag is mixed; the reaction slag is monitored in real time, and the slag after the completion of the reaction is obtained by controlling and simultaneously ensuring two parameters (a) and (b);
  • the temperature of the copper-containing reaction slag is 1080 ° C, the electric arc furnace is heated to raise the temperature to 1330 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 0.1, to the reaction melt Adding alkaline iron concentrate, converter sludge, alkaline pre-reduction pellets, high-calcium red mud after de-sodium, and increasing the alkalinity ratio of copper-containing reaction slag to 0.3; spraying natural gas, metal in slag The iron content is 1.6%;
  • Step 2 separation and recovery method 1:
  • the slag after the completion of the reaction is naturally settled, and the slag-gold is separated to obtain a molten copper-rich nickel phase, an iron-rich phase and a ferrosilicate-containing mineral phase, and at the same time, a zinc-containing component and a lead-containing component are formed. Enter the soot and recycle it as an oxide. Perform the following steps:
  • the iron-rich phase layer is poured into the heat preservation device and cooled as a raw material for direct reduction of nickel-iron;
  • the zinc-containing component, the lead-containing component, the indium-containing component and the cerium-containing component are volatilized, and are collected into the soot for recovery.
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 91%, and the zinc recovery rate is 92. %, the lead recovery rate is 92%, the sodium recovery rate is 95%, the potassium recovery rate is 96%, the nickel enrichment rate is 94%, the cobalt enrichment rate is 92%, and the gold enrichment rate is 91%.
  • the silver enrichment rate is 92%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: the cold nickel smelting slag obtained by the "osmosis process” process, the cold nickel smelting slag obtained by the “copper ice nickel blowing process” blowing slag through the depleting furnace is added to the submerged arc furnace, Limestone, SiO 2 , FeO and MgO are added to form a mixed slag; the mixed slag is heated to a molten state to form a copper-containing reaction slag, and an argon-nitrogen gas mixture having a preheating temperature of 900 ° C is sprayed, and the reaction is carried out.
  • the slag is mixed; the reaction slag is monitored in real time, and the slag after the completion of the reaction is obtained by controlling both parameters (a) and (b);
  • the temperature of the copper-containing reaction slag is 1320 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 0.8, both within the required range; using air, the spray size is ⁇ 150 ⁇ m Pulverized coal and natural gas, the content of metallic iron in the slag is 2.7%;
  • Step 2 separation and recovery method 4:
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 91%, the zinc recovery rate is 93%, and the lead recovery rate. It is 92%; the nickel enrichment rate is 93%, the cobalt enrichment rate is 95%, the gold enrichment rate is 91%, and the silver enrichment rate is 92%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: molten nickel smelting slag obtained from the slag smelting slag exiting process of the smelting process, and molten smelting slag obtained by the "copper ice nickel blowing" process slag slag passing through the slag outlet of the lean furnace Adding to the blast furnace, adding dolomite, red mud, MgO, using coke powder with a preheating temperature of 800 ° C and having an oxygen particle size of ⁇ 150 ⁇ m, and spraying the gas to heat the mixed slag to a molten state to form a copper-containing reaction slag. And the reaction slag is mixed; the reaction slag is monitored in real time, and the slag after the completion of the reaction is obtained by controlling and simultaneously ensuring two parameters (a) and (b);
  • the temperature of the copper-containing reaction slag is 1330 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 1.0, both within the required range; the metal iron content in the slag is 1.9 %;
  • reaction slag as a hot metallurgical flux, adjusting the composition of the copper-containing reaction slag to control the temperature of the copper-containing reaction slag;
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 93%, the zinc recovery rate is 92%, and the lead recovery rate. 94%, sodium recovery rate is 95%, potassium recovery rate is 96%, nickel enrichment rate is 91%, cobalt enrichment rate is 97%, gold enrichment rate is 92%, silver enrichment rate It is 93%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 slag mixing: molten nickel obtained by melting the smelting slag from the top blowing molten pool through the slag outlet of the settling electric furnace and molten nickel obtained by the "copper ice nickel blowing" process blowing slag through the slag outlet of the depleted furnace
  • the smelting slag is added to the side blowing furnace, and limestone is added at the same time to form mixed slag; the pulverized powder of ⁇ 150 ⁇ m is sprayed with air having a preheating temperature of 900 ° C, and the mixed slag is heated to a molten state to form a copper-containing reaction slag, and The reaction slag is mixed; the reaction slag is monitored in real time, and the slag after the completion of the reaction is obtained by controlling both parameters (a) and (b);
  • the temperature of the copper-containing reaction slag is 1340 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 1.2, both within the required range;
  • the metal iron content in the slag is 2.1 %;
  • Step 2 separation and recovery method 2:
  • the slag after the completion of the reaction is naturally settled, and the slag-gold is separated, and the copper-rich nickel phase and the middle-upper iron-rich phase and the iron-containing silicate mineral phase are obtained, and the zinc-containing component and the lead-containing component are simultaneously formed.
  • the following steps are carried out: (1) the iron-rich phase and the iron-containing silicate mineral phase are poured into the smelting device, and the middle and upper portions are melted by the method C in the separation and recovery method of the second step Slag pouring glass ceramics;
  • the finally obtained slag contains copper ⁇ 0.1%, iron recovery rate is 94%, zinc recovery rate is 92%, lead recovery rate is 92%, nickel enrichment rate is 97%, and cobalt enrichment rate is 93. %, the gold enrichment rate is 92%, and the silver enrichment rate is 92%.
  • a method for recovering valuable components from a nickel-containing smelting slag comprising the steps of:
  • Step 1 Slag mixing: molten nickel obtained by melting the smelting slag from the top-blown molten pool to obtain the molten nickel slag obtained by the slag discharge port of the sedimentation electric furnace, and the molten nickel obtained by the "copper ice nickel blowing" process blowing slag through the slag outlet of the depleted furnace
  • the smelting slag and the molten nickel smelting slag obtained by melting the smelting slag from the top blowing molten pool through the sedimentation electric furnace slag inlet are added to the thermal insulation pit, and limestone and Fe are added at the same time to form mixed slag; with oxygen-enriched air, the spray size is ⁇ 150 ⁇ m Bituminous coal, the mixed slag is heated to a molten state to form a copper-containing reaction slag, and the reaction slag is mixed; the reaction slag is monitored in real time, and the two parameters of (a) and (b) are simultaneously controlled to
  • the temperature of the copper-containing reaction slag is 1430 ° C;
  • the alkalinity CaO / SiO 2 ratio of the copper-containing reaction slag is 0.5, both within the required range;
  • the metal iron content in the slag is 0.8 %;
  • the slag contains copper ⁇ 0.1%, the iron recovery rate is 91%, the zinc recovery rate is 93%, and the lead recovery rate is 92. %; nickel enrichment rate is 96%, cobalt enrichment rate is 97%, gold enrichment rate is 91%, and silver enrichment rate is 92%.

Abstract

一种由含镍冶炼熔渣回收有价组分的方法，其包括将镍冶炼渣加入反应装置中，并加入钙系矿物与添加剂，形成混合熔渣，将混合熔渣加热至熔融状态作为反应熔渣，混合均匀，实时监测该反应熔渣，通过调控使混合后的反应熔渣同时满足条件a和条件b，获得反应后的熔渣；分离回收。该方法既可以充分利用熔融镍渣物理热资源和热态冶金熔剂，又可以处理冷态炉渣，通过加入添加剂，混合均匀，控制熔渣氧位，实现了熔渣冶金，实现镍冶炼熔渣中铜、铁同步分离技术，并解决目前炉渣大量堆积，环境污染问题，及重金属元素污染问题。

Description

一种由含镍冶炼熔渣回收有价组分的方法 技术领域
本发明属于熔渣冶金技术领域，具体涉及一种由含镍冶炼熔渣回收有价组分的方法。
背景技术
镍的火法冶炼过程中，产生大量镍冶炼渣，镍冶炼渣包括“造锍熔炼”工艺产生的镍熔炼渣、“铜冰镍吹炼”工艺吹炼后的贫化炉渣、顶吹熔炼产生的镍沉降炉渣。我国金川公司采用“闪速熔炼”与“顶吹熔炼”进行镍精矿冶炼，每年产生100万吨镍冶炼渣，目前累计堆积储存量超过1000万吨。镍冶炼渣中铜、铁、镍、钴、锌、铅、金、银等有价组分，长期堆放，既浪费资源，又污染环境。
镍冶炼渣中铁含量高达50wt％，远超目前26wt％可采品位，渣含铜高达0.2wt％以上，其中铁组分主要以铁橄榄石相存在。目前，镍冶炼渣的研究利用主要集中直接还原提铁或熔融炼铁。在还原过程中，铜、镍、钴组分还原进入铁水，不利于后续的炼钢过程。
镍冶炼渣不仅是重要的二次资源等，而且含有大量SiO2、CaO、MgO、Al2O3等冶金熔剂，化学反应活性强，是物理化学性质优良的熔渣体系，是热态冶金熟料，是成熟的冶金渣系。如何能有效地以镍冶炼熔渣为反应渣系，实现镍冶炼熔渣中铜、铁同步分离技术是目前市场所需。
发明内容
(一)要解决的技术问题
为了解决现有技术的上述问题，本发明一种由含镍冶炼熔渣回收有价组分的方法。本发明方法是一种新的熔渣冶金工艺，不仅降低渣含铜使渣含铜<0.1wt％，而且实现铜、镍、钴、铁、金、银、铅、锌、铟、铋、钠、钾等组分的高效回收，获得低铜含铁物料(铁精矿与生铁)，并解决目前炉渣大量堆积，环境污染问题，及重金属元素污染问题。
(二)技术方案
为了达到上述目的，本发明采用的主要技术方案包括：
一种由含镍冶炼熔渣回收有价组分的方法，其包括如下步骤：
S1、炉渣混合：将镍冶炼渣加入熔炼反应装置中，并加入钙系矿物与添加剂，形成混合熔渣，将混合熔渣加热至熔融状态作为反应熔渣，混合均匀，实时监测该反应熔渣，通过调控使混合后的反应熔渣同时满足条件a和条件b，获得反应后的熔渣，或将反应后的熔渣倒入保温装置；
其中，所述条件a为控制反应熔渣的温度为1100～1500℃；
所述条件b为控制反应熔渣的碱度CaO/SiO2比值＝0.15～1.5；
S2、分离回收：所述步骤S1反应后的熔渣，保温5～50min，沉降分离获得底部的富铜镍相、中下部的富铁相层与中上部的含铁硅酸盐矿物相层，同时生成含锌组分与含铅组分的烟灰，金组分、银组分进入富铜镍相，含镍组分、含钴组分分别进入富铜镍相与富铁相；对各相进行回收处理。
如上所述的方法，优选地，在所述步骤S1中，对于所述条件a调控的方法为：
当所述反应熔渣的温度＜1100℃时，利用反应装置自身的加热功能，或向熔渣中加入燃料与预热的氧化性空气，使反应熔渣的温度达到设定温度范围内，使反应 熔渣的温度达到1100～1500℃；
当所述反应熔渣的温度＞1500℃时，向反应熔渣中加入镍冶炼渣、高炉渣、铁合金渣、钢渣、冶金熔剂、含铁物料或含氟物料中的一种或几种，使混合熔渣的温度达到1100～1500℃；
对于所述条件b调控的方法为：
当所述反应熔渣中碱度CaO/SiO2比值<0.15时，向反应熔渣中加入碱性物料和/或碱性含铁物料；
当所述反应熔渣中碱度CaO/SiO2比值>1.5时，向反应熔渣中加入酸性物料和/或酸性含铁物料。
如上所述的方法，优选地，所述反应装置为保温装置或可转动的熔炼反应装置或带有渣口或铁口熔渣可流出的熔炼反应装置；其中，所述保温装置为可倾倒的熔炼反应渣灌或保温地坑；
所述可转动的熔炼反应装置为转炉、熔炼反应渣罐；
所述带有渣口或铁口熔渣可流出的熔炼反应装置为等离子炉、直流电弧炉、交流电弧炉、矿热炉、鼓风炉、高炉、感应炉、冲天炉、侧吹熔池熔炼炉、底吹熔池熔炼炉、顶吹熔池熔炼炉、反射炉、奥斯麦特炉、艾萨炉、瓦钮可夫熔池熔炼炉、侧吹回转炉、底吹回转炉、顶吹回转炉。
如上所述的方法，优选地，在所述步骤S1中，满足所述条件a和b的同时，应同时满足，控制所述反应熔渣中铜氧化物、镍氧化物、钴氧化物和铁氧化物分别还原为金属铜、金属镍、金属钴和FeO，熔渣中金属铁含量<3％。可通过加入还原剂、含碳的含铁物料中的一种或两种进行调控，其中，所述还原剂和/或含碳的含铁物料的用量为熔渣中铜氧化物、镍氧化物、钴氧化物和铁氧化物还原为金属铜、金属镍、金属钴和FeO的理论量110～140％；所述含碳的含铁物料为钢铁尘泥与烟灰、铁精矿含碳预还原球团、铁精矿含碳金属化球团、湿法炼锌挥发窑渣、焦炭炉尘泥与烟灰。
如上所述的方法，优选地，所述镍冶炼渣是“造锍熔炼”工艺产生的镍熔炼渣、铜冰镍吹炼”工艺吹炼后的吹炼渣、“铜冰镍吹炼”工艺吹炼后的吹炼渣经贫化产生的贫化炉渣、顶吹熔炼产生的镍熔炼渣、顶吹熔炼产生的镍熔炼渣经沉降产生的沉降渣中一种或多种；所述镍冶炼渣为熔融态或冷态，其中，熔融镍冶炼渣由镍冶炼炉出渣口获得，或将镍冶炼渣加热至熔融状态；
所述钙系矿物为石灰、石灰石、白云石、电石渣、赤泥或脱钠高钙后赤泥中的一种或几种；所述添加剂为SiO2、MgO、FeO、Fe2O3、MnO2、Al2O3、TiO2、Fe或Na2O中的一种或几种。
如上所述的方法，优选地，所述含铜物料为粗铜铜火法精炼渣、选铜尾矿、铜渣、锌冶炼渣、锌冶炼烟灰与尘泥、铅冶炼炉渣、铅锌尾矿、铅冰铜、砷冰铜、粗铅火法精炼渣、铅冶炼烟尘与烟灰、铅酸电池、铜冶炼烟灰与尘泥、杂铜、含铜垃圾、含铜电路板、锡冶炼渣、锡尾矿中的一种或几种；其中，所述锌冶炼渣为湿法炼锌与火法炼锌产生的锌冶炼渣，包括浸出渣、铁矾渣、铜镉渣、针铁矿渣、赤铁矿渣、挥发窑渣、竖罐炼锌渣、鼓风炉渣、烟化炉渣、旋涡炉渣、电炉炼锌渣，所述湿法炼锌渣在进炉前，需进行烘干、脱水处理；含铅冶炼渣为烟化炉炉渣与含铅熔炼渣，“ISP铅锌鼓风炉还原”或“烧结矿鼓风炉还原”或“固态高铅渣还原”或“液态高铅渣还原工艺”还原工艺产生含铅熔炼渣，含铅熔炼渣通过烟化炉冶炼产生含铅烟化炉渣；所述铜渣是含铜熔炼渣、含铜吹炼渣、贫化弃渣、浮选尾渣、湿法炼铜渣中的一种或多种，含铜熔炼渣产生于铜的火法冶炼工艺的“造锍熔炼”过程；所述含铜吹 炼渣产生于铜的火法冶炼工艺的“铜锍吹炼”过程；贫化弃渣为含铜熔炼渣与含铜吹炼渣贫化后弃渣，浮选尾渣为含铜熔炼渣与含铜吹炼渣选矿后尾渣。
如上所述的方法，优选地，所述燃料与还原剂为固体、液体或气体燃料中的一种或多种，以喷吹或投料的方式喷入，所述喷吹时，用0～1200℃的气体载入，所述气体是氧化性气体；
所述碱性物料为石灰粉、赤泥、脱钠后高钙赤泥、电石渣、白云石粉或生石灰粉中的一种或几种；所述碱性含铁物料为CaO/SiO2>1的碱性烧结矿、碱性铁精矿、碱性预还原球团、碱性金属化球团或钢渣中的一种或几种；
如上所述的方法，优选地，所述酸性物料为硅石、含金银硅石、粉煤灰、煤矸石中的一种或多种；所述酸性含铁物料为CaO/SiO2≤1的含铁物料、酸性烧结矿、酸性铁精矿、酸性预还原球团、酸性金属化球团、铜渣、含铅炉渣、锌冶炼渣、镍冶炼渣、铁合金渣、高炉渣中的一种或几种。
如上所述的方法，优选地，所述含铁物料是普通铁精矿、普通铁精矿直接还原铁，普通铁精矿烧结矿、普通铁精矿球团矿、普通铁精矿金属化球团、普通铁精矿含碳预还原球团、钢渣、锌冶炼渣、焦炭冶炼烟尘与尘泥、钢铁烟尘与尘泥、含镍冶炼渣、铜渣、铅冶炼渣、铜渣、锡冶炼渣、赤泥、脱钠后高钙赤泥、煤粉灰、硫酸烧渣中的一种或几种；所述钢铁烟尘与尘泥包括高炉瓦斯泥、转炉尘泥、电炉尘泥、热(冷)轧污泥、烧结粉尘、球团粉尘、出铁厂集尘、高炉瓦斯灰、电炉除尘灰、轧钢氧化铁皮；
所述含氟物料是萤石、CaF2或含氟高炉渣中的一种或几种；
所述湿法炼锌渣与尘泥需脱水、干燥。
进一步地，所述固体燃料与还原剂为煤粉、焦粉、焦炭、粉煤灰、烟煤或无烟煤中的一种或多种，形状为粒状或粉状，粒状物料粒度为5～25mm，粉状物料粒度为≤150μm；所述液体燃料与还原剂为重油，所述气体燃料与还原剂为煤气和/或天然气。
优选地，在所述步骤S1中，所述混合均匀为自然混合或搅拌混合，搅拌方式为氩气搅拌、氮气搅拌、氮气-氩气混合气搅拌、还原性气体搅拌、氧化性气体搅拌、电磁搅拌或机械搅拌中的一种或几种。
如上所述的方法，优选地，所述含铜物料、含铁物料和含氟物料均为球团或粉状物料或制粒；其中，粉状物料的粒度≤150μm，粒状物料粒度为5～25mm，粉状物料以喷吹的方式喷入，粒状物料以喷吹或投料的方式加入,所述喷吹所用的载入气体为预热的氩气、氮气、还原性气体(煤气和/或天然气)、氧化性气体中的一种或多种，所述预热的温度为0～1200℃。
其中，含铜物料与含铁物料为热态或冷态，其中热态物料由冶金炉出料口或出渣口直接获得。
如上所述的方法，优选地，在所述步骤S2中分离回收中，对于沉降在底部的所述富铜镍相、中部的富铁及上层的含铁硅酸盐矿物相可分别进行处理，或中上部的富铁与含铁硅酸盐矿物相结合处理，金组分与银组分迁移、富集进入富铜镍相，部分镍钴组分进入富铁相，含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收。
具体地，所述步骤S2中的分离回收，采用如下方法一到方法五中任一方法处理：
方法一、采用熔渣可流出熔炼反应装置时，反应完成后的熔渣分离后进行如下步骤：
S2-1-01、所述含铁硅酸盐矿物相，进行如下方法A-G中的任一种处理；
方法A：水淬或空冷后直接用作水泥原料；
方法B：部分或全部所述含铁硅酸盐矿物相返回到所述反应熔渣中作为热态冶金熔剂；
方法C：用于浇筑微晶玻璃或作为矿渣棉；方法D：所述含铁硅酸盐矿物相的熔渣氧化后空冷或水淬，方法包括：熔渣保留在熔炼反应装置内或将熔渣倒入保温装置，向含铁硅酸盐的熔渣中，吹入温度为0～1200℃的预热氧化性气体，并保证硅酸盐的熔渣温度>1450℃；当熔渣氧化亚铁的重量百分含量<1％，获得氧化后的熔渣；所述氧化后的熔渣直接空冷或水淬，用作矿渣水泥、水泥调整剂、水泥生产中的添加剂或水泥熟料；
进一步地，当含铁硅酸盐熔渣温度<1450℃，喷入预热燃料与预热的氧化性气体，燃烧放热、补充热量，或装置自身加热，使含铁硅酸盐熔渣温度>1450℃；
方法E：所述含铁硅酸盐矿物相用于生产高附加值的水泥熟料，包括如下步骤：
E-1、含铁硅酸盐矿物相保留在熔炼反应装置内或将熔渣倒入保温装置，向含铁硅酸盐矿物相的熔渣中，加入熔融钢渣、石灰、石灰石、铁合金炉渣、粉煤灰、碱性铁贫矿、铝土矿、熔融高炉渣、赤泥、脱钠后赤泥或电石渣中的一种或几种，充分混合，获得熔渣混合物料；
E-2、向上熔渣混合物料中吹入预热温度为0～1200℃的氧化性气体，并保证熔渣混合物料温度>1450℃；当氧化亚铁重量百分比含量<1％，获得氧化后的熔渣；
E-3、所述氧化后的熔渣，进行空冷或水淬，制得高附加值的水泥熟料；
方法F：所述含铁硅酸盐矿物相的熔渣作为高炉炼铁原料或直接还原炼铁原料：将含铁硅酸盐熔渣空冷、水淬或缓冷后，用作高炉炼铁或直接还原炼铁原料，直接还原后，采用磁选分离或电炉熔分，磁选产物为金属铁与尾矿，电炉熔分，产物为铁水与熔渣；
或将熔渣倒入保温装置后，采用熔渣改性后磁选分离，包括：向保温装置中的熔渣，吹入0～1200℃的预热的氧化性气体，并保证其熔渣温度>1250℃；
进一步地，当熔渣温度<1250℃，喷入预热燃料与预热的氧化性气体，燃烧放热、补充热量，或装置自身加热，使熔渣温度>1250℃；
将上述氧化后的熔渣缓冷至室温，破碎、磁选，产物为磁铁矿精矿与尾矿，尾矿作为建筑材料；
方法G：所述含铁硅酸盐矿物相进行还原炼铁，包括如下步骤：
G-1、含铁硅酸盐矿物相保留在熔炼反应装置内或将该熔渣倒入保温装置，或加入含铁物料，同时加入还原剂，进行熔融还原，实时监测反应熔渣，通过调控同时满足条件：反应熔渣的温度为1350～1670℃和反应熔渣的碱度CaO/SiO2比值＝0.6～2.4，获得反应完成后的熔渣；
其中，控制反应熔渣的温度的方法为：
当反应熔渣的温度＜1350℃，通过反应装置自身的加热，或向熔渣中加入燃料与预热的氧化性气体，使反应熔渣的温度达到1350～1670℃；
当反应熔渣的温度＞1670℃，向反应熔渣中加入冶金熔剂、含铁物料或含氟物料中的一种或几种，使反应熔渣的温度达到1350～1670℃，其中，所述冶金熔剂为含CaO或SiO2的矿物，具体为石英砂、含金银石英砂、赤泥、脱钠后高钙赤泥、电石渣、白云石或石灰石中的一种或几种；
控制反应熔渣的碱度的方法为：
当反应熔渣中碱度CaO/SiO2比值<0.6时，向熔渣中加入碱性物料和/或碱性含铁物料；
当反应熔渣中碱度CaO/SiO2比值>2.4时，向熔渣中加入酸性物料和/或酸性含铁物料；
G-2、所述G-1中熔融还原时还需向熔渣中喷吹0～1200℃预热后的氧化性气体进行熔融还原，形成还原后的熔渣；
G-3、分离回收：采用以下两种方法中的一种进行：
方法Ⅰ：将还原后的混合熔渣倒入保温渣罐，缓慢冷却至室温，获得缓冷渣；其中，金属铁沉降到反应装置的底部，形成铁坨，将剩余缓冷渣中含金属铁层，破碎至粒度20～400μm，磨矿，磁选分离出剩余金属铁与尾矿；
方法Ⅱ：还原后的混合熔渣，冷却沉降，渣-金分离，获得铁水与还原后的熔渣；所述还原后的熔渣，按照方法A～E中的一种或几种方法进行熔渣处理；所述铁水，送往转炉或电炉炼钢；
S2-1-02、所述富铜镍相，送往转炉或吹炼炉或作为铜钴镍分离的原料；
S2-1-03、部分所述含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收；
S2-1-04、部分含金组分与含银组分进入富铜镍相，镍、钴组分分别进入富铜镍相与富铁相；
S2-1-05、所述富铁相层进行水淬或空冷或倒入保温装置缓冷后或经人工分拣与重选结合获得，作为高炉炼镍铁原料或直接还原炼镍铁原料或熔融还原炼镍铁原料或浮选提铜原料；浮选过程中，浮选产物为铜精矿、镍精矿、含镍合金与铁精矿，铁精矿作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料；其中，所述直接还原过程采用转底炉、隧道窑、车底路、竖炉、回转窑或感应炉作为还原设备，利用气基或煤基还原技术，气基还原采用天然气和/或煤气，煤基还原采用无烟煤、烟煤、褐煤、焦煤、焦粉或焦炭中的一种或几种，控制还原温度为900～1400℃，控制碱度CaO/SiO2比值＝0.8～1.5；还原产生的煤气在熔渣表面二次燃烧，提供了热量，而且由炉内流出的煤气可以作为烘干炉料与保温装置的热源；
此外，因赤泥中含有钾、钠，尘泥与钢铁烟灰中含有铅、锌、铋、铟，所以添加这些原料时，部分铟组分、铋组分、含钾组分、含钠组分挥发，以氧化物形式进入烟尘。
方法二、采用熔渣可流出的熔炼反应装置时，获得的所述富铁相和含铁硅酸盐矿物相处理方法用方法一中所述方法A～G中一种或几种进行处理，倒入保温装置缓冷后的富铜镍相，送往转炉或吹炼炉或作为铜钴镍分离的原料。
方法三、采用熔渣可转动的转炉与反应渣罐时，获得熔融态所述含铁硅酸盐矿物相，处理方法用方法一中所述方法A～G中一种或几种进行处理；或含有所述富铁相采用方法一中步骤S2-1-05进行处理；所述熔融态或倒入保温装置缓冷后的富铜镍相，送往转炉或吹炼炉或作为铜钴镍分离的原料。
方法四、采用熔渣可转动的转炉与反应渣罐时，获得的熔融态所述含铁硅酸盐矿物相与富铁相，处理方法用方法一中所述方法A～G中一种或几种进行处理；所述熔融态富铜镍相，送往转炉或吹炼炉炼铜，或缓冷后碎磁选分离金属铁后再送往转炉或吹炼炉或作为铜钴镍分离的原料。方法五：采用保温装置时，或采用熔渣可流出的熔炼反应装置，将熔渣倒入保温装置时，进行如下步骤：
S201、沉降冷却：熔渣冷却至室温，获得缓冷渣；所述富镍铜相沉降到反应装置的底部，形成富铜镍坨；所述含铁硅酸盐矿物相上浮；中间为缓冷渣为富铁相，同时生成含锌组分与含铅组分；其中，金、银组分迁移到富铜镍相，镍、钴组分分别迁移到富铜镍相与富铁相；
S202、分离：人工取出沉降在底部的富铜镍坨，送往转炉或吹炼炉或作为铜钴镍分离的原料；所述富铁相作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料或熔融还原炼铁原料或浮选提铜原料；浮选产物为含铜精矿、镍精矿、镍合金与铁精矿，铁精矿作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料；
S203、人工取出上部的含铁硅酸盐矿物相，获得硅酸盐相作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料或水泥原料；
S204、部分锌组分与铅组分挥发，以氧化物形式进入烟尘回收；
S205、添加有赤泥中或尘泥与钢铁烟灰这些原料时，部分铟组分、铋组分、含钾组分、含钠组分挥发，进入烟尘回收。
如上所述方法，优选地，所述氧化性气体为预热的空气、氧气、富氧空气、氮气-空气、氩气-空气、氧气-氮气、氧气-氩气中的一种。
如上所述方法，优选地，在所述步骤S1中，所述混合均匀为自然混合与搅拌混合，所述搅拌混合的方式为氩气搅拌、氮气搅拌、氮气-氩气混合气搅拌、还原性气体搅拌、氧化性气体搅拌、电磁搅拌或机械搅拌中的一种或几种；
在所述步骤S2中，所述沉降为自然沉降或旋转沉降或离心沉降；进行冷却沉降时的冷却方式为自然冷却或旋转冷却或离心冷却，所述分离时，用重力分选法是摇床分选、溜槽分选或者二者相结合。
与现有技术相比，本发明的特点是：
(1)本发明的由含镍冶炼熔渣回收有价组分的方法，既可以处理热态熔渣，充分利用熔融镍冶炼渣物理热资源和热态冶金熔剂，又可以处理冷态炉渣，通过调整熔渣物理化学性质，利用含镍熔渣成熟的物理化学性质，实现了含镍熔渣冶金工艺；并解决目前炉渣大量堆积，环境污染问题，及重金属元素污染问题；
(2)熔渣中的熔渣冶金反应，加入有钙系矿物，使铁橄榄石解体，铁氧化物充分释放出来，形成游离态的铁氧化物，实现富铁相长大与沉降，熔渣中的含铁组分聚集、长大与沉降，富铁相包括金属铁、FeO相、铁橄榄石相中的多种，作为高炉炼铁或直接还原或熔融还原炼铁的原料；
同时钙系矿物有效改造粘度，是粘度降低，有助于含铜组分沉降，
(3)混合熔渣中的铜组分、镍组分、钴组分、金银组分分别迁移、富集于富铜镍相，并实现长大与沉降，其中，富铜镍相包括铜、白冰铜、铜冰镍、富铜镍钴相、含铁组分中的多种，或部分铜镍钴组分进入富铁相；
(4)混合熔渣中的锌组分、铅组分分别迁移、富集于烟灰中，并实现回收；部分含铟组分、铋组分、含钾组分、含钠组分挥发进入烟尘进行回收；
(6)分离沉降在不同部位的富铜镍相、铁氧化物，实现熔渣中铜组分、铁组分的高效回收，渣含铜<0.1wt％，可以处理固态含铜物料；
(7)熔渣实现调质，可作为水泥原料或建筑材料或代替碎石作骨料和路材；
(8)本发明方法加入添加剂，一是用于减小粘度，二是用于降低熔点，在一定温度(1100～1450℃)下有助于富铜镍相沉降，使沉降分离后获得富铁相为低铜富铁相与含铁硅酸盐相，其中富铁相与含铁硅酸盐相的含铜量小于0.1％，可以通过直接还原或熔融还原炼铁，获得铁精矿与金属铁；
(9)本发明方法可连续或间断的进行，满足了工业生产的需要。
(三)有益效果
本发明的有益效果是：
(1)本发明的原料可以是冷态炉渣，冷态炉渣处理，不仅可以实现炉渣中铜组分、 镍组分、钴组分、金组分、银组分、铋组分、钠组分、钾组分、铁组分、锌组分、铅组分有价组分的综合利用，可以解决目前炉渣大量堆积，环境污染问题。
(2)本发明的原料可以是出渣口中流出的液态熔融镍冶炼渣，蕴含着丰富的热能资源，具有高温度、高热量的特点，充分利用了熔渣物理热资源，高效节约能源；液态熔融铜渣含有大量的热态冶金熔剂，是物理化学性质优良的熔渣体系，实现了熔渣冶金。
(3)本发明通过调整熔渣物理化学性质，喷吹气体，控制氧势，使熔渣中铜组分、镍组分、钴组分、金银组分迁移、富集到富铜镍相，实现聚集、长大与沉降。
(4)本发明方法中，加入冷态物料与熔融镍冶炼渣避免了熔渣温度过高，提高保温装置的寿命；加入冷态物料与镍冶炼渣提高了原料处理量，不仅可以处理液态熔渣，而且可以处理少量冷态物料，原料适应性强；加入冷态物料实现了反应释放的化学热与熔渣物理热的高效利用。
(5)本发明通过调整熔渣物理化学性质，控制氧势，同时加入钙系矿物，使橄榄石相中铁氧化物释放出来，富集于富铁相，实现聚集、长大与沉降，富铁相包括金属铁、FeO相、铁橄榄石相中的多种，作为高炉炼铁或直接还原或熔融还原炼铁的原料；熔渣中锌组分、铅组分、铟组分、铋组分、钠组分、钾组分挥发，进入烟尘加以回收。
(6)本发明方法自然冷却过程中，添加剂的加入，熔渣中铜组分、镍组分、钴组分、金银组分富集于富铜镍相，并实现聚集、长大与沉降，富铜镍相包括有铜、白冰铜、铜冰镍、铜镍钴相、含铁组分中的多种，或部分铜镍钴组分进入富铁相，熔渣中铁组分富集于富铁相，并实现聚集、长大与沉降，装有熔渣的保温装置置于旋转平台上旋转，加速富铜镍相、富铁相的聚集、长大与沉降；含氟物料的加入，加速富铜镍相、富铁相的长大与沉降。
(7)本发明方法熔渣中铁氧化物几乎消失，矿物可磨性增加，熔渣实现调质。
(8)本发明方法采用人工分拣、磁选、重选或渣-金分离的方法，分别对分布在上部、中部与底部的低铁硅酸盐矿物相、富铁相、富铜镍相进行分离，实现熔渣中铜组分、镍组分、钴组分、金银组分、铟组分、铋组分、铁组分的高效回收，获得的含铁硅酸盐相与富铁组分的渣含铜<0.1wt％，由于富铜镍相、富铁相沉降在中、下部，因此，需分选炉渣量小，矿物可磨性增加，磨矿、磁选与重选成本低；后续的分离过程采用物理选矿(磁选或重选)，分离的介质为水，分离过程中不会产生环境污染，熔渣处理工艺具有流程短、操作简单、回收率高，具有高效、清洁、环保的特点；尾矿作为水泥原料、建筑材料、代替碎石作骨料、路材使用。
(9)本发明通过富铜镍相沉降，含铁硅酸盐相与富铁组分的含铜量小于0.1％，作为高炉炼铁或直接还原或熔融还原的原料,获得金属铁与铁水。
(10)本发明既充分利用熔融渣物理热资源和热态冶金熔剂，又可以处理冷态炉渣，实现了熔渣冶金，熔渣中铜组分、镍组分、钴组分、金银组分富集于富铜镍相，并实现聚集、长大与沉降，铁组分富集于富铁相，实现聚集、长大与沉降，采用人工分拣、磁选、重选与渣-金沉降结合的方法，分离沉降在不同部位的富铜镍相和富铁相，实现熔渣中铜组分、镍、钴、金银组分、铁组分的高效回收；可以处理固态含铜物料，该方法工艺流程短、金属回收率高、生产成本低、原料适应性强、处理量大、环境友好、经济收益高、可解决冶金资源与热能高效回收利用问题。
具体实施方式
一种由含镍冶炼熔渣回收有价组分的方法，具体包括以下步骤：
步骤1，炉渣混合：
将含镍冶炼熔渣，加入保温装置或熔渣可流出的熔炼反应装置中并加入钙系矿物与添加剂，形成混合熔渣；
将混合熔渣加热至熔融状态，形成反应熔渣；混合均匀，实时监测含镍的反应熔渣，通过调控同时保证如下(a)和(b)两个参数，获得反应完成后的熔渣，或将反应完成后的熔渣倒入保温装置；
(a)反应熔渣的温度为1100～1500℃；
(b)反应熔渣的碱度CaO/SiO2比值＝0.15～1.5；
调控方法为：
对应(a)：
控制反应熔渣的温度在设定温度范围的方法为：
当反应熔渣的温度＜设定温度范围下限时，通过反应装置自身的加热功能，或反应熔渣中加入燃料与预热的氧化性空气，使反应熔渣的温度达到设定温度范围内；
当反应熔渣的温度＞设定温度范围上限时，向反应熔渣中加入镍冶炼渣、高炉渣、铁合金渣、钢渣、含铜物料、冶金熔剂、含铁物料或含氟物料中的一种或几种，使混合熔渣的温度达到设定温度范围内；
对应(b)：
当反应熔渣中碱度CaO/SiO2比值<0.15时，向反应熔渣中加入碱性物料或碱性含铁物料中的一种或几种；
当反应熔渣中碱度CaO/SiO2比值>1.5时，向反应熔渣中加入酸性物料或酸性含铁物料中的一种或几种；
步骤2，分离回收：
保温5～50min，反应完成后的熔渣，沉降，渣-金分离，获得底部熔融态的富铜镍、中下部熔融态的富铁相与中上部熔融态的含铁硅酸盐矿物相，同时生成含锌组分与含铅组分，其中，铟、铋、镍、钴、金、银组分迁移到所述富镍铜相，部分镍钴组分迁移到富铁相；
采用以下方法中的一种对各相进行回收处理：
方法一：采用熔渣可流出熔炼反应装置时，反应完成后的熔渣进行如下步骤：
(1)熔融态含铁硅酸盐矿物相，进行熔渣处理；
(2)熔融态富铜，送往转炉或吹炼炉或作为铜钴镍分离的原料；
(3)部分铅组分、锌组分、铟组分、铋组分、钠组分、钾组分挥发进入烟尘回收；
(4)富铁相进行水淬或空冷或倒入保温装置缓冷或经人工分拣与重选结合获得，作为高炉炼铁原料或直接还原炼镍铁原料或熔融还原炼镍铁原料或浮选提铜镍原料；浮选产物为铜精矿、镍精矿、含镍合金与铁精矿，铁精矿作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料；所述直接还原过程采用转底炉、隧道窑、车底路、竖炉、回转窑或感应炉作为还原设备，利用气基或煤基还原技术，气基还原为天然气和/或煤气，煤基还原为无烟煤、烟煤、褐煤、焦煤、焦粉或焦炭中的一种或几种，还原温度为900～1400℃，碱度CaO/SiO2比值＝0.8～1.5。
其中，步骤(1)中的含铁硅酸盐矿物相，进行熔渣处理，采用方法A～G中的一种：
方法A：含铁硅酸盐矿物相作为水泥原料：
含铁硅酸盐矿物相水淬或空冷直接作为水泥原料或进一步处理成高附加值的水泥原料。
方法B：部分或全部含铁硅酸盐矿物相返回到所述含铜反应熔渣：
部分或全部含铁硅酸盐矿物相返回到含铜反应熔渣，作为热态冶金熔剂，调整含铜反应熔渣成分，控制含铜反应熔渣温度。
方法C：含铁硅酸盐矿物相浇筑微晶玻璃或作为矿渣棉。
方法D：含铁硅酸盐熔渣氧化后空冷或水淬：
(1)向熔炼反应装置内的含铁硅酸盐熔渣中，吹入预热的氧化性气体，当熔渣氧化亚铁含量<1％，完成熔渣的氧化，获得氧化后的熔渣，其中，氧化性气体的预热温度为0～1200℃；并在整个过程中，保证(c)硅酸盐熔渣温度>1450℃；
对应(c)采用的控制方法：
当含铁硅酸盐熔渣温度<1450℃，喷入预热燃料与预热的氧化性气体，燃烧放热、补充热量，或装置自身加热，使硅酸盐熔渣温度>1450℃；
(2)氧化后的熔渣直接空冷或水淬，用作矿渣水泥、水泥调整剂、水泥生产中的添加剂或水泥熟料。
方法E：含铁硅酸盐熔渣处理生产高附加值的水泥熟料：
(1)向熔炼反应装置内的含铁硅酸盐熔渣中，加入熔融钢渣、石灰、石灰石、铁合金炉渣、粉煤灰、碱性铁贫矿、铝土矿、熔融高炉渣、赤泥、脱钠后赤泥或电石渣中的一种或几种，充分混合，获得熔渣混合物料；
(2)向熔渣混合物料中吹入预热的氧化性气体，当氧化亚铁含量<1％，完成熔渣的氧化，获得氧化后的熔渣，其中，氧化性气体的预热温度为0～1190℃；并在整个过程中，保证(d)熔渣混合物料温度>1450℃；温度控制方法同方法D步骤(1)中的硅酸盐熔渣温度控制方法；
(3)氧化后的熔渣，进行空冷或水淬，制得高附加值的水泥熟料。
方法F：所述含铁硅酸盐矿物相熔渣作为高炉炼铁原料或直接还原炼铁原料：将含铁硅酸盐矿物相的熔渣空冷、水淬或缓冷后，用作高炉炼铁或直接还原炼铁原料，直接还原后，采用磁选分离或电炉熔分，磁选产物为金属铁与尾矿，电炉熔分，产物为铁水与熔渣；
或将所述含铁硅酸盐矿物相的熔渣倒入保温装置后，采用以下方法进行分离：熔渣改性后磁选分离：向保温装置中的熔渣，吹入0～1200℃的预热的氧化性气体，并保证其熔渣温度>1250℃，完成熔渣中磁铁矿的转化；将上述氧化后的熔渣缓冷至室温，破碎、磁选，产物为磁铁矿精矿与尾矿，尾矿作为建筑材料。
方法G：含铁硅酸盐熔渣熔融还原炼铁：
(1)将含铁硅酸盐熔渣保留在熔炼反应装置内或将熔渣倒入保温装置，向熔融态熔渣加入含铁物料，还原剂，进行熔融还原，实时监测反应熔渣，通过调控同时保证如下(a)和(b)两个参数，获得反应完成后的熔渣；
(a)反应熔渣的温度为1350～1670℃；
(b)反应熔渣的碱度CaO/SiO2比值＝0.6～2.4；
调控方法为：
对应(a)：
控制反应熔渣的温度在设定温度范围的方法为：
当反应熔渣的温度＜设定温度范围下限1350℃时，通过反应装置自身的加热功能，或向熔渣中加入燃料与预热的氧化性气体，使反应熔渣的温度达到设定温度范围1350～1670℃内；
当反应熔渣的温度＞设定温度范围上限1350℃时，向反应熔渣中加入冶金熔剂、含铁物料或含氟物料中的一种或几种，使反应熔渣的温度达到设定温度范围 1350～1670℃内；
对应(b)：
当反应熔渣中碱度CaO/SiO2比值<0.6时，向熔渣中加入碱性物料和/或碱性含铁物料；
当反应熔渣中碱度CaO/SiO2比值>2.4时，向熔渣中加入酸性物料和/或酸性含铁物料；
(2)向熔渣中喷吹预热后的氧化性气体进行熔融还原，形成还原后的熔渣，其中：氧化性气体的预热温度为0～1200℃，并在喷吹过程中，通过调控同时保证(a)和(b)两个参数：
(a)反应完成后的熔渣的温度为1350～1670℃；
(b)反应完成后的熔渣的碱度CaO/SiO2比值＝0.6～2.4；
其中，设定温度范围和碱度调控方法同方法G步骤(1)；
(3)分离回收：
采用以下方法中的一种：
方法Ⅰ：进行如下步骤：
(a)冷却：将还原后的混合熔渣倒入保温渣罐，冷却至室温，获得缓冷渣；
(b)分离：金属铁沉降到反应装置的底部，形成铁坨，人工取出铁坨；将剩余缓冷渣中含金属铁层，破碎至粒度20～400μm，磨矿，磁选分离出剩余金属铁与尾矿；
(c)尾矿的回收利用，作为水泥原料、建筑材料、代替碎石作骨料、路材或磷肥使用；
方法Ⅱ：进行如下步骤：
(a)还原后的混合熔渣，冷却沉降，渣-金分离，获得铁水与还原后的熔渣；
(b)还原后的熔渣，进行炉外熔渣处理，具体方式为：采用步骤2的分离回收方法一中的方法A～E中的一种或几种，进行熔渣处理；
(c)铁水，送往转炉或电炉炼钢；
(d)含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收；
(e)部分铟组分、铋组分、钠组分、钾组分挥发进入烟尘；
(f)还原产生的煤气在熔渣表面二次燃烧，提供了热量，而且由炉内流出的煤气可以作为烘干炉料与保温装置的热源。
方法二：采用熔渣可流出的熔炼反应装置时，反应完成后的熔渣进行如下步骤：
(1)冷态富铜镍相，送往转炉炼铜或吹炼炉或作为铜钴镍分离的原料；
(2)富铁相层与含铁硅酸盐矿物相采用处理方法用方法一中所述方法A～G中一种或几种进行处理；
(3)部分含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收；
(4)部分铟组分、铋组分、钠组分、钾组分挥发进入烟尘。
方法三：采用熔渣可转动的转炉与反应渣罐时，反应完成后的熔渣进行如下步骤：
(1)熔融态含铁硅酸盐矿物相，进行熔渣处理，具体处理方式为：采用步骤2的分离回收方法一中的方法A～G中的一种或几种进行熔渣处理；或含有所述富铁相采用方法一中的步骤(4)进行处理；
(2)熔融态富铜镍相或倒入保温装置缓冷后，送往转炉或吹炼炉或作为铜钴镍分离的原料；
(3)部分铅组分、锌组分、铟组分、铋组分、钠组分、钾组分挥发进入烟尘；
方法四：采用熔渣可转动的转炉与反应渣罐时，反应完成后的熔渣进行如下步骤：
(1)熔融态含铁硅酸盐矿物相与富铁相，进行熔渣处理，具体方式为：采用步骤2的分离回收方法一中的方法A～G中的一种或几种进行处理；
(2)熔融态富铜镍相，送往转炉或吹炼炉或作为铜钴镍分离的原料；
(3)部分铅组分、锌组分、铟组分、铋组分、钠组分、钾组分挥发进入烟尘；
方法五：采用保温装置时，或采用熔渣可流出的熔炼反应装置，将熔渣倒入保温装置时，反应完成后的熔渣进行如下步骤：
(1)沉降冷却：反应完成后的熔渣冷却至室温，获得缓冷渣；富铜镍相沉降到反应装置的底部，形成富铜镍坨；含铁硅酸盐矿物相上浮；富铜镍相和含铁硅酸盐矿物中间的缓冷渣为富铁相，同时生成含锌组分与含铅组分；金银组分迁移到富铜镍相；
(2)分离：人工取出沉降在底部的富铜镍坨，往转炉或吹炼炉或作为铜钴镍分离的原料；中部的富铁相层作为高炉炼镍铁原料或直接还原炼镍铁原料或熔融还原炼镍铁原料或浮选提铜原料；在浮选过程中，浮选产物为铜精矿、镍精矿、镍合金与铁精矿，铁精矿作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料；
所述直接还原过程采用转底炉、隧道窑、车底路、竖炉、回转窑或感应炉作为还原设备，利用气基或煤基还原技术，气基还原采用天然气和/或煤气，煤基还原采用无烟煤、烟煤、褐煤、焦煤、焦粉或焦炭中的一种或几种，控制还原温度为900～1400℃，控制碱度CaO/SiO2比值＝0.8～1.5；
(3)人工取出上部的含铁硅酸盐矿物相，作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料或作为水泥原料、建筑材料、代替碎石作骨料、路材使用；
(4)部分铅组分、锌组分、铟组分、铋组分、钠组分、钾组分挥发进入烟尘。
如上所述的方法中，优选地，所述的步骤1与2中，镍冶炼渣为熔融态或冷态，其中：熔融镍冶炼渣由镍冶炼炉出渣口获得，或将镍冶炼渣加热至熔融状态；镍冶炼渣是“造锍熔炼”工艺产生的镍熔炼渣、铜冰镍吹炼”工艺吹炼后的吹炼渣、“铜冰镍吹炼”工艺吹炼后的吹炼渣经贫化产生的贫化炉渣、顶吹熔炼产生的镍熔炼渣、顶吹熔炼产生的镍熔炼渣经沉降产生的沉降渣中一种或多种；
所述的步骤1与2中铜物料是粗铜火法精炼渣、铜渣、选铜尾矿、锌冶炼渣、锌冶炼烟灰与尘泥、铅冶炼炉渣、铅锌尾矿、铅冰铜、砷冰铜、粗铅火法精炼渣、铅冶炼烟灰与尘泥、铅酸电池、铜冶炼烟灰与尘泥、杂铜、含铜垃圾、含铜电路板、锡冶炼渣、锡尾矿中的一种或几种；锌冶炼渣为湿法炼锌与火法炼锌产生的锌冶炼渣，包括浸出渣、铁矾渣、铜镉渣、针铁矿渣、赤铁矿渣、挥发窑渣、竖罐炼锌渣、鼓风炉渣、烟化炉渣、旋涡炉渣、电炉炼锌渣；含铅炉渣为烟化炉炉渣与含铅熔炼渣，“ISP铅锌鼓风炉还原”或“烧结矿鼓风炉还原”或“固态高铅渣还原”或“液态高铅渣还原工艺”还原工艺产生含铅熔炼渣，含铅熔炼渣通过烟化炉冶炼产生含铅烟化炉渣；
所述的步骤1与2中，铜渣是含铜熔炼渣、含铜吹炼渣、贫化弃渣、浮选尾渣、湿法炼铜渣中的一种或多种，含铜熔炼渣产生于铜的火法冶炼工艺的“造锍熔炼”过程；含铜吹炼渣产生于铜的火法冶炼工艺的“铜锍吹炼”过程；贫化弃渣为含铜熔炼渣与含铜吹炼渣贫化后弃渣，浮选尾渣为含铜熔炼渣与含铜吹炼渣选矿后尾渣。
如上所述的方法中，优选地，所述的步骤1与2中，熔渣可流出的熔炼反应装置为可转动的熔炼反应装置或带有渣口或铁口的熔炼反应装置；其中：
所述的保温装置为可倾倒的熔炼反应渣灌、保温地坑；
所述的可转动的熔炼反应装置为转炉、熔炼反应渣罐；
所述的带有渣口或铁口熔渣可流出的熔炼反应装置为等离子炉、直流电弧炉、交流电弧炉、矿热炉、鼓风炉、高炉、感应炉、冲天炉、侧吹熔池熔炼炉、底吹熔池熔炼炉、顶吹熔池熔炼炉、反射炉、奥斯麦特炉、艾萨炉、瓦钮可夫熔池熔炼炉、侧吹回转炉、底吹回转炉、顶吹回转炉。
如上所述的方法中，优选地，所述的步骤1中，钙系矿物具体为石灰、石灰石、白云石、电石渣、赤泥或脱钠后高钙赤泥中的一种或几种；添加剂为SiO2、MgO、FeO、Fe2O3、MnO2、Al2O3、TiO2、Fe或Na2O中的一种或几种。
所述的步骤1中，通过调控同时保证如下(a)和(b)两个参数，同时保证熔渣中铜氧化物、镍氧化物、钴氧化物和铁氧化物分别还原为金属铜、金属镍、金属钴和FeO，熔渣中金属铁含量<3％。可通过加入还原剂、含碳的含铁物料中的一种或两种进行调控，其中，所述还原剂和/或含碳的含铁物料的用量为熔渣中铜氧化物、镍氧化物、钴氧化物和铁氧化物还原为金属铜、金属镍、金属钴和FeO的理论量110～140％；所述含碳的含铁物料为钢铁尘泥与烟灰、铁精矿含碳预还原球团、铁精矿含碳金属化球团、湿法炼锌挥发窑渣、焦炭炉尘泥与烟灰。
所述的步骤1与2中，所述的燃料与还原剂为固体、液体或气体燃料中的一种或多种，以喷吹或投料的方式喷入，载入气体为预热的氧化性气体、氮气或氩气中的一种或多种，预热温度为0～1200℃；固体燃料与还原剂为煤粉、焦粉、焦炭、粉煤灰、烟煤或无烟煤中的一种或多种，形状为粒状或粉状，粒状物料粒度为5～25mm，粉状物料粒度为≤150μm，液体燃料与还原剂为重油，气体燃料与还原剂为煤气和/或天然气；
所述的步骤1与2中冶金熔剂为含CaO或SiO2的矿物，具体为石英砂、含金银石英砂、赤泥、脱钠后高钙赤泥、电石渣、白云石或石灰石中的一种或几种；
所述的1与2中含铁物料是普通铁精矿、普通铁精矿直接还原铁，普通铁精矿烧结矿、普通铁精矿球团矿、普通铁精矿金属化球团、普通铁精矿含碳预还原球团、钢渣、锌冶炼渣、焦炭冶炼烟尘与尘泥、钢铁烟尘与尘泥、含镍冶炼渣、铜渣、铅冶炼渣、锌冶炼渣、锡冶炼渣、赤泥、脱钠后高钙赤泥、煤粉灰、硫酸烧渣中的一种或几种；所述钢铁烟尘与尘泥包括高炉瓦斯泥、转炉尘泥、电炉尘泥、热(冷)轧污泥、烧结粉尘、球团粉尘、出铁厂集尘、高炉瓦斯灰、电炉除尘灰、轧钢氧化铁皮。
所述含铜物料与含铁物料为热态或冷态，其中热态物料由冶金炉出料口或出渣口直接获得。
所述湿法炼锌渣与尘泥需经脱水、干燥。
在上述的原料中，锌冶炼渣与烟灰、铅冶炼渣与烟灰含有铟与铋、铅、银、锌、铋；赤泥中含有钠与钾，钢铁烟尘与尘泥含有铟、铋、银、钠与钾，以上物料都有铁，铅冶炼渣与锌冶炼渣都含有铜，铜烟灰与尘泥含有铟与铋，因此在发明的方法中，铟、铋、钠、钾、锌、铅会以氧化物的形式进入烟尘，从而进行回收。
所述的1与2中含氟物料是萤石、CaF2或含氟高炉渣中的一种或几种；
所述的步骤1与2中，含铜物料、含铁物料和含氟物料均为球团或粉状物料或制粒；其中，粉状物料的粒度≤150μm，粒状物料粒度为5～25mm，粉状物料以喷吹的方式喷入，粒状物料以喷吹或投料的方式加入，载入气体为预热的氩气、氮气、还原性气体(煤气和/或天然气)、氧化性气体中的一种或多种，预热温度为0～1200℃；
所述的步骤1与2中采用喷吹方式进行加料，具体采用耐火喷枪插入熔渣或置 于反应熔渣上部或侧面或底部吹入中的一种或几种。
所述的步骤1与2中，熔渣反应过程中，熔渣中铜组分、镍组分、钴组分、金银组分富集于富铜镍相，并实现聚集、长大与沉降，铁组分富集于富铁相，实现聚集、长大与沉降，熔渣中锌组分、铅组分、铟组分、铋组分分别进入烟尘回收。所述的控制混合熔渣的温度在设定温度范围的方法中：
当混合熔渣的温度＞设定温度上限时，加入镍冶炼渣、含铜物料、冶金熔剂、含铁物料或含氟物料中的一种或几种，目的是避免温度过高，保护耐火材料；加入含氟物料的另一个作用是降低粘度，加速熔渣中富铜镍相、富铁相的聚集、长大与沉降。
所述步骤1与2中，调整碱度时，所述的碱性物料为石灰粉、赤泥、脱钠后赤泥、电石渣、白云石粉或生石灰粉中的一种或几种；所述的碱性含铁物料为CaO/SiO2>1的含铁物料，所述的碱性含铁物料为碱性烧结矿、碱性铁精矿、碱性预还原球团、碱性金属化球团、钢渣或高炉渣中的一种或几种。
所述步骤1与2中，调整碱度时，酸性物料为硅石、粉煤灰、煤矸石中的一种或多种；酸性含铁物料为CaO/SiO2≤1的含铁物料，酸性含铁物料为酸性烧结矿、酸性铁精矿、酸性预还原球团、酸性金属化球团、铜渣、铅冶炼渣、锌冶炼渣、镍冶炼渣、锡冶炼渣、铁合金渣、高炉渣中的一种或几种；
在步骤1中，保证(a)和(b)两个参数的同时，要使熔渣充分混合，混合方式为自然混合或搅拌混合，搅拌方式为以下方式中的一种：氩气搅拌、氮气搅拌、氮气-氩气混合气搅拌、还原性气体搅拌、氧化性气体、电磁搅拌、机械搅拌中的一种或多种，气体预热温度为0～1200℃；
在步骤1与2中，熔渣中富铜镍相、富铁相聚集、长大与沉降，有利于硅酸盐上浮，其中，富铜镍相包括铜、白冰铜、铜冰镍、富铜镍钴相、含铁组分中的多种，或部分铜镍钴组分进入富铁相，富铁相包括金属铁、FeO相、铁橄榄石相中的多种；
所述的步骤1与2中，氧化性气体为预热的空气、氧气或富氧空气中的一种，喷吹方式为采用耐火喷枪插入熔渣或置于反应熔渣上部或侧面或底部吹入中的一种或几种，预热温度为0～1200℃；
所述的步骤2中，直接还原过程采用转底炉、隧道窑、车底路、竖炉、回转窑或感应炉作为还原设备，利用气基或煤基还原技术，气基还原为天然气和/或煤气，煤基还原为无烟煤、烟煤、褐煤、焦煤、焦粉或焦炭中的一种或几种，还原温度为900～1400℃，碱度CaO/SiO2比值＝0.8～1.5；
所述的步骤1与2中，含铜物料与含铁物料为热态或冷态，所述的热态物料是从冶金炉中直接产出的热态物料，热态物料温度为200～1750℃；
在步骤2中，需要冷却沉降时，其中冷却方式为自然冷却或旋转冷却或离心冷却，沉降方式为自然沉降或旋转沉降或离心沉降；
进一步地，旋转与离心的具体操作为：装有反应完成后的熔渣的装置置于旋转平台上，按照一定速度进行旋转，旋转速度依熔渣质量与保温装置高度或深度而定，旋转时间依熔渣质量与熔渣凝固情况而定；将装有反应完成后的熔渣的装置置于旋转平台上旋转，目的是加速富铜镍相、富铁相聚集、长大与沉降，有利于硅酸盐上浮，缩短沉降时间，改善沉降效果，提高生产效率；
所述的步骤2中，反应完成后的熔渣冷却过程中，由于密度不同与矿物大小不同，大部分富铜镍相、富铁相沉降于中下部；
所述的步骤2中，反应完成后的熔渣中铜组分、镍组分、钴组分、金银组分继 续富集于富铜镍相，实现长大与沉降，或部分富集于富铁相；混合熔渣中铁组分继续富集于富铁相，并实现长大与沉降。
所述的步骤2中，分离时，可采用重力分选法是摇床分选、溜槽分选或者二者相结合。
如采用本发明的方法，最后获得的富铁相及含铁硅酸盐相的渣中含铜≤0.1％，最后铁的回收率为≥91％，锌的回收率为≥92％，铅的回收率为≥92％，镍的富集率为≥91％，钴的富集率为≥91％，金的富集率为≥90％，银的富集率为≥90％。
为了更好的解释本发明，以便于理解，通过具体实施方式，对本发明作详细描述。其中，以下实施例中所用检测方法与原料未明确指出的，均可采用本领域常规技术，除非另有说明，本发明中所用的百分数均为重量百分数。
以下实施例1～10中的步骤(1)熔渣混合时，通过调控保证的(a)和(b)两个参数具体为：
(a)含铜与铁的反应熔渣的温度为1100～1500℃；
(b)含铜与铁的反应熔渣碱度CaO/SiO2比值＝0.15～1.5。
实施例1
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：
将由“造锍工艺”工艺熔炼炉出渣口获得的熔融镍冶炼渣加入直流电弧炉，同时加入石灰，以及SiO2、MgO、Al2O3，形成混合熔渣；将混合熔渣加热至熔融状态，形成反应熔渣，并使反应熔渣实现自然混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；对应(a)：反应熔渣的温度为1660℃，采用耐火喷枪插入反应熔渣中，以氮气为载入气，喷入常温粉状粒度≤150μm的铜渣、含铜烟灰、杂铜和含铜垃圾和含铜电路板，同时加入高炉瓦斯泥、电炉尘泥、转炉尘泥、普通铁精矿直接还原铁和高炉瓦斯灰，使温度降至1490℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值＝2.6，向反应熔渣中加入硅石、粉煤灰和煤矸石混合物，使含铜反应熔渣碱度比值降至0.28；熔渣中金属铁含量为1.1％；
步骤2，分离回收采用方法一：
保温45min，反应完成后的熔渣自沉降，渣-金分离，获得熔融态富铜镍层、富铁层与含铁硅酸盐矿物相，同时生成锌组分与铅组分进入烟尘，进行如下步骤：(1)熔融态含铁硅酸盐矿物相，进行炉外熔渣处理，采用方法F，硅酸盐熔渣空冷后，用作直接还原炼铁原料，直接还原过程中，采用回转窑作为还原设备，利用气基还原技术，气基还原剂为天然气和煤气，还原温度为950℃，碱度CaO/SiO2比值为0.8，还原后采用电炉熔分温度为1550℃，产物为金属铁水与熔渣；(2)熔融态富铜镍相，送往转炉；(3)富铁相层倒入保温装置，空冷后作为高炉炼铁原料；(4)锌组分、铟组分、铅组分、铋组分、钾组分、钠组分挥发，以氧化物形式进入烟尘回收；渣含铜<0.1％，锌回收率为92％，铅回收率为93％，铟回收率为92％，铋回收率为94％，钠回收率为95％，钾回收率为96％，铁回收率为96％，镍的富集率为93％，钴的富集率为95％，金的富集率为93％，银的富集率为94％。其中，在本发明的所有实施例中，渣含铜是指富铜镍相分离后的渣相，具体为富铁相与硅酸盐矿物相中的含铜量，镍、钴的富集率是指在富铜镍相中镍、钴的含量占原料中对应镍、钴总量的百分比，金、银的富集率是指富铜镍相中金、银的含量占原料中金、银总量的百分比。
实施例2
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：
将由“铜冰镍吹炼”工艺吹炼炉出渣口获得的熔融镍冶炼渣加入可倾倒的熔炼反应渣灌，同时加入石灰石、白云石、赤泥以及FeO和Fe2O3，形成混合熔渣；用预热温度为600℃的富氧空气粒度为20mm无烟煤与焦粒，并喷吹天然气，将混合熔渣加热至熔融状态，形成含铜反应熔渣；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)含铜反应熔渣温度为1660℃，采用耐火喷枪插入反应熔渣中，以氩气为载气，喷入常温粉状粒度≤150μm铜渣、镍冶炼渣、含铜烟灰、杂铜和、钢铁烧结粉尘、烧结球团粉尘、出铁厂粉尘、普通铁精矿直接还原铁，使温度降至1480℃；(b)含铜反应熔渣碱度CaO/SiO2比值为2.4，向反应熔渣中加入酸性铁精矿、酸性预还原球团、含铅熔炼渣、含铅烟化炉渣的混合物，使含铜反应熔渣碱度比值降至1.2；熔渣中金属铁含量为3％；
步骤2，分离回收采用方法二：
保温50min，反应完成后的熔渣自然沉降，渣-金分离，获得熔融态富铜镍相层、富铁相层与硅酸盐矿物相，同时生成锌组分、铅组分与铟组分，进入烟尘，以氧化物形式加以回收，进行如下步骤：
(1)熔融含铁硅酸盐矿物相与富铁相，采用方法G进行炉外熔渣处理，熔渣熔融还原炼铁，具体步骤如下：
(1-1)熔渣倒入可倾倒的转炉中，向熔渣中加入粒度为20mm无烟煤与烟煤，进行熔融还原，实时监测反应熔渣，通过调控同时保证如下(a)反应熔渣的温度为1350～1670℃，和(b)反应熔渣的碱度CaO/SiO2比值＝0.6～2.4两个参数，获得反应完成后的熔渣；
对应(a)：反应熔渣的温度为1460℃，在温度范围内；
对应(b)：反应熔渣中碱度CaO/SiO2比值为0.8时，在碱度范围内；
(1-2)向熔渣中喷吹预热200℃的氧化性气体(富氧空气)进行熔融还原，形成还原后的混合熔渣，并在喷吹过程中，通过调控同时保证(a)反应熔渣的温度为1350～1670℃，和(b)反应熔渣的碱度CaO/SiO2比值＝0.6～2.4两个参数，
(1-3)分离回收：
(a)还原后的混合熔渣，自然沉降，获得铁水与还原后的熔渣；
(b)还原后的熔渣，采用步骤2方法一中方法A处理做成高附加值水泥原料；
(c)铁水，送往转炉或电炉炼钢；
(d)含锌组分、含铅组分、铋组分与铟组分挥发，以氧化物形式进入烟尘回收；
(e)含钠组分、含钾组分挥发，进入烟尘回收；
(2)熔融态富铜镍相，送往转炉。
(3)锌组分与铅组分挥发，以氧化物形式进入烟尘回收。
锌回收率为95％，铅回收率为93％，铟回收率为96％，铋回收率为96％，钠回收率为97％，钾回收率为98％，渣含铜<0.1％，铁回收率为97％；镍的富集率为92％，钴的富集率为96％，金的富集率为90％，银的富集率为90％。
实施例3
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由“铜冰镍吹炼”工艺吹炼渣经贫化炉出渣口获得的熔融镍冶炼渣，加入反射炉，同时加入石灰石与脱钠后高钙赤泥，形成混合熔渣；用预热温度为900℃的氧气，喷吹粒度为20mm无烟煤、焦粒与煤粉，将混合熔渣加热至熔 融状态，形成含铜反应熔渣，机械搅拌混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1685℃，向反应熔渣中加入酸性金属化球团、铜渣，同时加入含铜烟灰、含铅炉渣、普通铁精矿球团矿、轧钢氧化铁鳞和普通铁精矿含碳预还原球团，使温度降至1420℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值＝2.3，向反应熔渣中加入石英砂、赤泥、锌冶炼渣的混合物，使含铜反应熔渣的碱度比值降至1.2；熔渣中金属铁含量为1.8％；
步骤2，分离回收采用方法二：
保温35min，反应完成后的熔渣自然沉降，渣-金分离，获得熔融态富铜镍相、富铁相层与硅酸盐矿物相，同时生成含锌组分与含铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：
(1)熔融态富铜镍相，送往转炉；
(2)熔融态富铁相层与硅酸盐矿物相作为直接还原炼镍铁原料；还原过程中，锌组分、铅组分、铋组分与铟组分挥发，进入烟尘；直接还原过程中，采用转底炉，还原温度为1200℃，碱度CaO/SiO2比值＝1.0，粒度为≤150μm的无烟煤与煤粉；
(3)含锌组分、铟组分、铋组分与含铅组分挥发，进入烟尘回收，渣含铜<0.1％，铁的回收率为96％，锌的回收率为93％，铅的回收率为92％，铟回收率为93％，铋回收率为94％；镍的富集率为93％，钴的富集率为97％，金的富集率为91％，银的富集率为92％。
实施例4
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将顶吹熔池熔炼炉出渣口获得的熔融镍冶炼渣加入等离子炉，同时加入白云石、MgO、Al2O3、以及Fe，形成混合熔渣；将混合熔渣加热至熔融状态，形成含铜反应熔渣，并使反应熔渣电磁搅拌，实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)含铜反应熔渣的温度为1670℃，向反应熔渣中加入赤泥、硫酸烧渣、萤石、铅冰铜、含铅烟灰、含锌烟灰、砷冰铜和湿法炼锌渣，使温度降至1440℃；(b)含铜反应熔渣的碱度CaO/SiO2比值为2.0，向反应熔渣中加入铜渣，使含铜反应熔渣碱度比值降至0.7；用空气，喷吹天然气、粒度为20mm的焦粒，熔渣中金属铁含量为0.9％；
步骤2，分离回收采用方法二：
保温14min，反应完成后的熔渣自然沉降，渣-金分离，获得熔融态富铜镍相、富铁相层与含铁硅酸盐矿物相，同时生成含锌组分与含铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：
(1)熔融态富铜镍相，送往转炉；
(2)熔融态富铁相层与含铁硅酸盐矿物相步骤2分离回收方法一中方法F，氧化改性磁选分离：①将熔渣倒入保温渣罐，向熔渣中喷入预热温度为900℃的富氧空气，实现磁铁矿的转化；②缓冷至室温，磁选分离，获得铁精矿与尾矿；
(3)部分含锌组分、铋组分、铟组分与含铅组分挥发，以氧化物形式进入烟尘进行回收，渣含铜<0.1％，铁的回收率为95％，锌的回收率为92％，铅的回收率为92％，铟回收率为93％，铋回收率为94％；镍的富集率为93％，钴的富集率为94％，金的富集率为92％，银的富集率为94％。
实施例5
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由顶吹熔池熔炼产生熔炼渣经沉降电炉出渣口获得的熔融镍冶炼渣加入保温渣罐，同时加入石灰石以及Fe，形成混合熔渣；用预热温度为600℃的富氧空气，喷吹粒度≤150μm烟煤，将混合熔渣加热至熔融状态，形成含铜反应熔渣，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1430℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为1.5，均在要求范围内；熔渣中金属铁含量为2.2％；
步骤2，分离回收采用方法五：
将反应完成后的熔渣倒入保温渣罐中，保温25min,进行炉外熔渣处理，进行如下步骤：
(1)沉降冷却：反应完成后的熔渣旋转冷却至室温，获得缓冷渣；富铜镍相沉降到反应装置的底部，形成富铜镍坨；含铁硅酸盐矿物相上浮；富铜镍相金属坨和硅酸盐矿物中间缓冷渣为富铁相，同时生成含锌组分与含铅组分；
(2)分离：人工取出沉降在底部的富铜镍坨，产物送往转炉；中部的富铁相层直接送往转炉炼铁；
(3)人工取出上部的含铁硅酸盐矿物相，获得硅酸盐尾矿，作为水泥原料使用；
(4)部分含锌组分与含铅挥发，以氧化物形式进入烟尘回收，渣含铜<0.1％，铁的回收率为92％，锌的回收率为94％，铅的回收率为95％；镍的富集率为95％，钴的富集率为93％，金的富集率为91％，银的富集率为92％。
实施例6
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由“造锍工艺”工艺熔炼炉出渣口获得的熔融镍冶炼渣、由“铜冰镍吹炼”工艺吹炼炉出渣口获得的熔融镍冶炼渣加入交流电弧炉，同时加入石灰、MgO、Al2O3、Fe2O3，形成混合熔渣；将混合熔渣加热至熔融状态，形成含铜反应熔渣，喷入预热温度为800℃的氩气，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1080℃，电弧炉加热，使温度升至1330℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为0.1，向反应熔渣中加入碱性铁精矿、转炉污泥、碱性预还原球团、脱钠后高钙赤泥，使含铜反应熔渣的碱度比值升至0.3；喷入天然气，熔渣中金属铁含量为1.6％；
步骤2，分离回收采用方法一：
保温38min，将反应完成后的熔渣自然沉降，渣-金分离，获得熔融态富铜镍相、富铁相和含铁硅酸盐矿物相，同时生成含锌组分与含铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：
(1)含铁硅酸盐矿物相采用步骤2分离回收方法一中方法A，水淬直接作为水泥原料；
(2)熔融态富铜镍相，送往转炉；
(3)富铁相层倒入保温装置冷却后作为直接还原炼镍铁的原料；
(4)含锌组分、含铅组分、含铟组分与含铋组分挥发，进入烟尘进行回收，渣含铜<0.1％，铁的回收率为91％，锌的回收率为92％，铅的回收率为92％，钠回收率为95％，钾回收率为96％，镍的富集率为94％，钴的富集率为92％，金的富集率为91％，银的富集率为92％。
实施例7
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由“造锍工艺”工艺获得的冷态镍冶炼渣、由“铜冰镍吹炼”工艺吹炼渣经贫化炉获得的冷态镍冶炼渣加入矿热炉，同时加入石灰石、SiO2、FeO以及MgO，形成混合熔渣；将混合熔渣加热至熔融状态，形成含铜反应熔渣，喷吹预热温度为900℃的氩气-氮气混合气，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1320℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为0.8，均在要求范围内；采用空气，喷吹粒度≤150μm煤粉与天然气，熔渣中金属铁含量为2.7％；
步骤2，分离回收采用方法四：
保温36min，将反应完成后的熔渣自然沉降，渣-金分离，获得熔融态富铜镍相、富铁相与含铁硅酸盐矿物相，同时生成锌组分与铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：
(1)熔融态富铜镍相，送往转炉；
(2)熔融态富铁相层与含铁硅酸盐矿物具体采用步骤2的分离回收方法一中的方法F，水淬后，作为直接还原炼镍铁的原料；
(3)部分含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收，渣含铜<0.1％，铁的回收率为91％，锌的回收率为93％，铅的回收率为92％；镍的富集率为93％，钴的富集率为95％，金的富集率为91％，银的富集率为92％。
实施例8
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由“造锍工艺”工艺熔炼炉出渣口获得的熔融镍冶炼渣、由“铜冰镍吹炼”工艺吹炼渣经贫化炉出渣口获得的熔融镍冶炼渣加入鼓风炉，同时加入白云石、赤泥、MgO，采用预热温度为800℃的氧气粒度≤150μm的焦粉，并喷入煤气，将混合熔渣加热至熔融状态，形成含铜反应熔渣，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1330℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为1.0，均在要求范围内；熔渣中金属铁含量为1.9％；
步骤2，分离回收采用方法三：
保温19min，将反应完成后的熔渣自然沉降，渣-金分离，获得富铜镍相与中上部的含铁硅酸盐矿物相和富铁相，同时生成含锌组分与含铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：
(1)熔融态含铁硅酸盐矿物相，倒入熔炼装置，进行炉外熔渣处理，具体采用步骤2的分离回收方法一中的方法B，将中上部的熔渣全部返回到含铜反应熔渣，作为热态冶金熔剂，调整含铜反应熔渣成分，控制含铜反应熔渣温度；
(2)熔融态富铜镍相，送往转炉或吹炼炉；
(3)部分含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收，渣含铜<0.1％，铁的回收率为93％，锌的回收率为92％，铅的回收率为94％，钠回收率为95％，钾回收率为96％，镍的富集率为91％，钴的富集率为97％，金的富集率为92％，银的富集率为93％。
实施例9
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：由顶吹熔池熔炼产生熔炼渣经沉降电炉出渣口获得的熔融镍冶炼渣与“铜冰镍吹炼”工艺吹炼渣经贫化炉出渣口获得的熔融镍冶炼渣加入侧吹炉，同时加入石灰石，形成混合熔渣；采用预热温度为900℃的空气喷入≤150μm焦粉，将混合熔渣加热至熔融状态，形成含铜反应熔渣，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1340℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为1.2，均在要求范围内；熔渣中金属铁含量为2.1％；
步骤2，分离回收采用方法二：
保温31min，将反应完成后的熔渣自然沉降，渣-金分离，获得富铜镍相与中上部的富铁相和含铁硅酸盐矿物相，同时生成含锌组分与含铅组分，进入烟尘，以氧化物形式回收，进行如下步骤：(1)富铁相和含铁硅酸盐矿物相倒入熔炼装置，采用步骤2的分离回收方法一中的方法C，将中上部熔渣浇筑微晶玻璃；
(2)熔融态富铜镍相，缓冷后用步骤二中的方法四进行处理；(3)部分含锌组分与含铅组分，以氧化物形式进入烟尘；
最后获得的渣含铜<0.1％，铁的回收率为94％，锌的回收率为92％，铅的回收率为92％；镍的富集率为97％，钴的富集率为93％，金的富集率为92％，银的富集率为92％。
实施例10
一种由含镍冶炼熔渣回收有价组分的方法，包括以下步骤：
步骤1，炉渣混合：将由顶吹熔池熔炼产生熔炼渣经沉降电炉出渣口获得的熔融镍冶炼渣、“铜冰镍吹炼”工艺吹炼渣经贫化炉出渣口获得的熔融镍冶炼渣与由顶吹熔池熔炼产生熔炼渣经沉降电炉出渣口获得的熔融镍冶炼渣加入保温地坑，同时加入石灰石以及Fe，形成混合熔渣；用富氧空气，喷吹粒度≤150μm烟煤，将混合熔渣加热至熔融状态，形成含铜反应熔渣，并使反应熔渣实现混合；实时监测反应熔渣，通过调控同时保证(a)和(b)两个参数，获得反应完成后的熔渣；
对应(a)：含铜反应熔渣的温度为1430℃；(b)：含铜反应熔渣的碱度CaO/SiO2比值为0.5，均在要求范围内；熔渣中金属铁含量为0.8％；
步骤2，分离回收采用方法五：
将反应完成后的熔渣进行如下步骤：
(1)沉降冷却：保温48min，反应完成后的熔渣冷却至室温，获得缓冷渣；富铜镍相沉降到反应装置的底部，形成富铜坨；含铁硅酸盐矿物相上浮；富铜镍相和硅酸盐矿物中间缓冷渣为富铁相，同时生成含锌组分与含铅组分；
(2)分离：人工取出沉降在底部的富铜镍坨；中部的富铁相层直接还原后，磁选分离金属镍铁；
(3)人工取出上部的硅酸盐矿物相，作为水泥原料使用；
(4)部分锌组分与铅组分挥发，以氧化物形式进入烟尘回收，渣含铜<0.1％，铁的回收率为91％，锌的回收率为93％，铅的回收率为92％；镍的富集率为96％，钴的富集率为97％，金的富集率为91％，银的富集率为92％。
以上所述，仅是本发明的较佳实施例而已，并非是对本发明做其它形式的限制，任何熟悉本专业的技术人员可能利用上述揭示的技术内容加以变更或改型为等同变化的等效实施例。但是凡是未脱离本发明技术方案内容，依据本发明的技术实质对以上实施例所作的任何简单修改、等同变化与改型，仍属于本发明技术方案的保护范围。

Claims (10)

  1. 一种由含镍冶炼熔渣回收有价组分的方法，其特征在于，其包括如下步骤：
    S1、炉渣混合：将镍冶炼渣加入反应装置中，并加入钙系矿物与添加剂，形成混合熔渣，将混合熔渣加热至熔融状态作为反应熔渣，混合均匀，实时监测该反应熔渣，通过调控使混合后的反应熔渣同时满足条件a和条件b，获得反应后的熔渣，或将反应后的熔渣倒入保温装置；
    其中，所述条件a为控制的熔渣温度在1100～1500℃；
    所述条件b为控制熔渣的碱度CaO/SiO2比值＝0.15～1.5；
    S2、分离回收：所述步骤S1反应后的熔渣，保温5～50min，沉降分离，获得底部的富铜镍相、中下部的富铁相与中上部的含铁硅酸盐矿物相，同时生成含锌组分与含铅组分的烟尘，金组分、银组分迁移、富集进入富铜镍相，含镍组分、含钴组分迁移进入富铜镍相与富铁相；对各相进行回收处理。
  2. 如权利要求1所述的方法，其特征在于，在所述步骤S1中，所述条件a调控的方法为：
    当所述反应熔渣的温度＜1100℃，利用反应装置自身的加热，或向熔渣中加入燃料与预热的氧化性空气，使反应熔渣的温度达到设定温度范围内，使反应熔渣的温度达到1100～1500℃；
    当所述反应熔渣的温度＞1500℃，向反应熔渣中加入镍冶炼渣、含铜物料、高炉渣、钢渣、铁合金渣、冶金熔剂、含铁物料或含氟物料中的一种或几种，混合均匀，使混合的反应熔渣的温度达到1100～1500℃；
    所述条件b调控的方法为：
    当所述反应熔渣中碱度CaO/SiO2比值<0.15时，向反应熔渣中加入碱性物料和/或碱性含铁物料；
    当所述反应熔渣中碱度CaO/SiO2比值>1.5时，向反应熔渣中加入酸性物料和/或酸性含铁物料。
  3. 如权利要求1所述的方法，其特征在于，所述反应装置为保温装置或可转动的熔炼反应装置或带有渣口或铁口熔渣可流出的熔炼反应装置；其中，所述保温装置为可倾倒的熔炼反应渣灌或保温地坑；
    所述可转动的熔炼反应装置为转炉、熔炼反应渣罐；
    所述带有渣口或铁口熔渣可流出的熔炼反应装置为等离子炉、直流电弧炉、交流电弧炉、矿热炉、鼓风炉、高炉、感应炉、冲天炉、侧吹熔池熔炼炉、底吹熔池熔炼炉、顶吹熔池熔炼炉、反射炉、奥斯麦特炉、艾萨炉、瓦钮可夫熔池熔炼炉、侧吹回转炉、底吹回转炉、顶吹回转炉。
  4. 如权利要求1所述的方法，其特征在于，在所述步骤S1中，满足所述条件a和b的同时，应同时满足，控制所述反应熔渣中铜氧化物、镍氧化物、钴氧化物和铁氧化物分别还原为金属铜、金属镍、金属钴和FeO，熔渣中金属铁含量<3％。
  5. 如权利要求1所述的方法，其特征在于，所述镍冶炼渣是“造锍熔炼”工艺产生的镍熔炼渣、“铜冰镍吹炼”工艺吹炼后的吹炼渣、“铜冰镍吹炼”工艺吹炼后的吹炼渣经贫化产生的贫化炉渣、顶吹熔炼产生的镍熔炼渣、顶吹熔炼产生的镍熔炼渣经电炉沉降产生的沉降渣中一种或多种；所述镍冶炼渣为熔融态或冷态，其中：熔融 镍冶炼渣由镍冶炼炉出渣口获得，或将镍冶炼渣加热至熔融状态；
    所述钙系矿物为石灰、石灰石、白云石、电石渣、赤泥或脱钠后高钙赤泥中的一种或几种；
    所述添加剂为SiO2、MgO、FeO、Fe2O3、MnO2、Al2O3、TiO2、Fe或Na2O中的一种或几种。
  6. 如权利要求2所述的方法，其特征在于，所述含铜物料为粗铜火法精炼渣、选铜尾矿、铜渣、锌冶炼渣、锌冶炼烟灰与尘泥、铅锌尾矿、铅冶炼渣、铅冰铜、砷冰铜、粗铅火法精炼渣、铅冶炼烟灰与尘泥、铅酸电池、铜冶炼烟灰与尘泥、杂铜、含铜垃圾、含铜电路板、锡冶炼渣、锡尾矿中的一种或几种；其中，所述锌冶炼渣为湿法炼锌与火法炼锌产生的锌冶炼渣，包括浸出渣、铁矾渣、铜镉渣、针铁矿渣、赤铁矿渣、挥发窑渣、鼓风炉渣、烟化炉渣、旋涡炉渣、竖罐炼锌渣、电炉炼锌渣；含铅冶炼渣为烟化炉炉渣与含铅熔炼渣，“ISP铅锌鼓风炉还原”或“烧结矿鼓风炉还原”或“固态高铅渣还原”或“液态高铅渣还原工艺”还原工艺产生含铅熔炼渣，含铅熔炼渣通过烟化炉冶炼产生含铅烟化炉渣；所述铜渣是含铜熔炼渣、含铜吹炼渣、贫化弃渣、浮选尾渣、湿法炼铜渣中的一种或多种，含铜熔炼渣产生于铜的火法冶炼工艺的“造锍熔炼”过程；所述含铜吹炼渣产生于铜的火法冶炼工艺的“铜锍吹炼”过程；贫化弃渣为含铜熔炼渣与含铜吹炼渣贫化后弃渣，浮选尾渣为含铜熔炼渣与含铜吹炼渣选矿后尾渣；
    所述含铁物料为普通铁精矿、普通铁精矿直接还原铁，普通铁精矿烧结矿、普通铁精矿球团矿、普通铁精矿金属化球团、普通铁精矿含碳预还原球团、钢渣、锌冶炼渣、焦炭冶炼烟尘与尘泥、钢铁烟尘与尘泥、含镍冶炼渣、铜渣、铅冶炼渣、锌冶炼渣、锡冶炼渣、赤泥、脱钠后高钙赤泥、煤粉灰、硫酸烧渣中的一种或几种；所述钢铁烟尘与尘泥包括高炉瓦斯泥、转炉尘泥、电炉尘泥、热或冷轧污泥、烧结粉尘、球团粉尘、出铁厂集尘、高炉瓦斯灰、电炉除尘灰、轧钢氧化铁皮；
    所述含铜物料与含铁物料为热态或冷态，其中热态物料由冶金炉出料口或出渣口直接获得；所述冶金熔剂为含CaO或SiO2的矿物与炉渣；所述含氟物料为萤石、CaF2或含氟高炉渣中的一种或几种；
    湿法炼锌渣、湿法炼铜渣与尘泥需经脱水、干燥；
    所述燃料为固体、液体或气体燃料中的一种或多种，所述燃料以喷吹或投料的方式加入，所述喷吹时，采用载入气体为预热的氧化性气体，所述预热的温度为0～1200℃；
    所述碱性物料为石灰粉、赤泥、脱钠后高钙赤泥、电石渣、白云石粉或生石灰粉中的一种或几种；所述碱性含铁物料为CaO/SiO2＞1的碱性烧结矿、碱性铁精矿、碱性预还原球团、碱性金属化球团、钢渣或高炉渣中的一种或几种；所述酸性物料为硅石、含金银硅石、粉煤灰、煤矸石中的一种或多种；所述酸性含铁物料为CaO/SiO2≤1的含铁物料、酸性烧结矿、酸性铁精矿、酸性预还原球团、酸性金属化球团、铜渣、铅冶炼渣、锌冶炼渣、镍冶炼渣、锡冶炼渣、铁合金渣、高炉渣中的一种或几种。
  7. 如权利要求2所述的方法，其特征在于，所述含铜物料、含铁物料和含氟物料均为球团或粉状物料或制粒；其中，粉状物料的粒度≤150μm，粒状物料粒度为5～25mm，粉状物料以喷吹的方式喷入，粒状物料以喷吹或投料的方式加入，所述喷 吹所用的载入气体为预热的氩气、氮气、还原性气体、氧化性气体中的一种或多种，所述预热的温度为0～1200℃。
  8. 如权利要求1-7中任一项所述的方法，其特征在于，所述步骤S2分离回收中，进行如下处理：
    含有热态或冷态所述富铜镍相，送往转炉或吹炼炉或作为铜钴镍分离的原料；
    所述含锌组分与含铅组分挥发，以氧化物形式进入烟尘回收；
    含镍组分、含钴组分迁移进入富铜镍相与富铁相；
    含有所述含铁硅酸盐矿物相和/或所述富铁相，进行如下方法A-G中的任一种处理；
    方法A：水淬或空冷后，直接作为水泥原料；
    方法B：部分或全部所述熔融态返回到所述反应熔渣中作为热态冶金熔剂；
    方法C：用于浇筑微晶玻璃或作为矿渣棉；
    方法D：所述含铁硅酸盐矿物相和/或所述富铁相保留在熔炼反应装置内或将其倒入保温装置，向熔渣中，吹入温度为0～1200℃的预热氧化性气体，并保证熔渣温度>1450℃；当熔渣氧化亚铁的重量百分含量<1％，获得氧化后的熔渣；所述氧化后的熔渣直接空冷或水淬，用作矿渣水泥、水泥调整剂、水泥生产中的添加剂或水泥熟料；
    方法E：用于生产高附加值的水泥熟料：
    E-1、将所述含铁硅酸盐矿物相和/或所述富铁相保留在熔炼反应装置内或将熔渣倒入保温装置，向熔渣中，加入熔融钢渣、石灰、石灰石、铁合金炉渣、粉煤灰、碱性铁贫矿、铝土矿、熔融高炉渣、赤泥、脱钠后赤泥或电石渣中的一种或几种，充分混合，获得熔渣混合物料；
    E-2、向上熔渣混合物料中吹入预热温度为0～1190℃的氧化性气体，并保证熔渣混合物料温度>1450℃；当氧化亚铁重量百分比含量<1％，获得氧化后的熔渣；
    E-3、对所述氧化后的熔渣，进行空冷或水淬，制得高附加值的水泥熟料；
    方法F：所述含铁硅酸盐矿物相和/或所述富铁相熔渣作为高炉炼铁原料或直接还原炼铁原料：将含铁硅酸盐矿物相和/或所述富铁相的熔渣空冷、水淬或缓冷后，用作高炉炼铁或直接还原炼铁原料，直接还原后，采用磁选分离或电炉熔分，磁选产物为金属铁与尾矿，电炉熔分，产物为铁水与熔渣；
    或将所述含铁硅酸盐矿物相和/或所述富铁相的熔渣倒入保温装置后，采用以下方法进行分离：熔渣改性后磁选分离：向保温装置中的熔渣，吹入0～1200℃的预热的氧化性气体，并保证其熔渣温度>1250℃，完成熔渣中磁铁矿的转化；将上述氧化后的熔渣缓冷至室温，破碎、磁选，产物为磁铁矿精矿与尾矿，尾矿作为建筑材料；
    方法G：所述含铁熔渣进行还原炼铁，包括如下步骤：
    G-1、将所述含铁硅酸盐矿物相和/或所述富铁相的熔渣保留在熔炼反应装置内或将熔渣倒入保温装置，向熔融态熔渣中加入含铁物料、还原剂，进行熔融还原，实时监测反应熔渣，通过调控同时满足条件：反应熔渣的温度为1350～1670℃和反应熔渣的碱度CaO/SiO2比值＝0.6～2.4，获得反应完成后的熔渣；
    其中，控制反应熔渣的温度的方法为：
    当反应熔渣的温度＜1350℃，通过反应装置自身的加热，或向熔渣中加入燃料与预热的氧化性气体，使反应熔渣的温度达到1350～1670℃；
    当反应熔渣的温度＞1670℃，向反应熔渣中加入冶金熔剂、含铁物料或含氟物料中的一种或几种，使反应熔渣的温度达到1350～1670℃；其中，所述冶金熔剂为含CaO或SiO2的矿物；
    控制反应熔渣的碱度的方法为：
    当反应熔渣中碱度CaO/SiO2比值<0.6时，向熔渣中加入碱性物料和/或碱性含铁物料；
    当反应熔渣中碱度CaO/SiO2比值>2.4时，向熔渣中加入酸性物料和/或酸性含铁物料；
    G-2、所述G-1中熔融还原时还需向熔渣中喷吹0～1200℃预热后的氧化性气体进行熔融还原，形成还原后的熔渣；
    G-3、分离回收：采用以下方法中的一种：
    方法Ⅰ：将还原后的混合熔渣倒入保温渣罐，冷却至室温，获得缓冷渣；其中，金属铁沉降到反应装置的底部，形成铁坨，将剩余缓冷渣中含金属铁层，破碎至粒度20～400μm，磨矿，磁选分离出剩余金属铁与尾矿；
    方法Ⅱ：还原后的混合熔渣，冷却沉降，渣-金分离，获得铁水与还原后的熔渣；所述还原后的熔渣，按照方法A～E中的一种或几种方法进行熔渣处理；所述铁水，送往转炉或电炉炼钢；
    或含有所述富铁相进行水淬或空冷或倒入保温装置缓冷或经人工分拣与重选结合获得，作为高炉炼镍铁原料或直接还原炼镍铁原料或熔融还原炼镍铁原料或浮选提铜镍原料；浮选产物为含铜精矿、镍精矿、镍铁合金与铁精矿，铁精矿作为高炉炼铁原料或直接还原炼铁原料或熔融还原炼铁原料；所述直接还原过程采用转底炉、隧道窑、车底路、竖炉、回转窑或感应炉作为还原设备，利用气基或煤基还原，气基还原采用天然气和/或煤气，煤基还原采用无烟煤、烟煤、褐煤、焦煤、焦粉或焦炭中的一种或几种，控制还原温度为900～1400℃，控制碱度CaO/SiO2比值＝0.8～1.5。
  9. 如权利要求8的方法，其特征在于，所述氧化性气体为预热的空气、氧气、富氧空气、氮气-空气、氩气-空气、氧气-氮气、氧气-氩气中的一种，所述预热的温度为0～1200℃；所述燃料与还原剂为固体、液体或气体燃料中的一种或多种，以喷吹或投料的方式喷入，所述喷吹载入气体为预热的氧化性气体，所述预热的温度为0～1200℃；所述固体燃料与还原剂为煤粉、焦粉、焦炭、粉煤灰、烟煤或无烟煤中的一种或多种，形状为粒状或粉状，粒状物料粒度为5～25mm，粉状物料粒度为≤150μm，所述液体燃料与还原剂为重油，所述气体燃料与还原剂为煤气和/或天然气。
  10. 如权利要求1所述的方法，其特征在于，在所述步骤S1中，所述混合均匀为自然混合或搅拌混合，搅拌方式为氩气搅拌、氮气搅拌、氮气-氩气混合气搅拌、还原性气体搅拌、氧化性气体搅拌、电磁搅拌或机械搅拌中的一种或几种；
    在所述步骤S2中，所述沉降为自然沉降或旋转沉降或离心沉降；进行冷却沉降时的冷却方式为自然冷却或旋转冷却或离心冷却，所述分离时，用重力分选法是摇床分选、溜槽分选或者二者相结合。
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Families Citing this family (7)

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Publication number Priority date Publication date Assignee Title
CN108728659B (zh) * 2018-06-20 2020-08-11 中国恩菲工程技术有限公司 镍渣贫化方法
CN111778407A (zh) * 2020-05-22 2020-10-16 金川集团股份有限公司 一种卡尔多炉吹炼含硫粗铜炉渣的处理方法
CN111778408A (zh) * 2020-05-22 2020-10-16 金川集团股份有限公司 一种用直流电弧炉处理自热炉炉渣生产合金的方法
CN115141937A (zh) * 2021-03-29 2022-10-04 东北大学 一种铜镍冶炼熔渣混合贫化及铁组分长大的方法
CN114480863B (zh) * 2022-04-18 2022-07-22 中国恩菲工程技术有限公司 金属镍渣的资源化利用方法
CN114990348A (zh) * 2022-05-30 2022-09-02 金川镍钴研究设计院有限责任公司 一种铜镍渣生产粗铜的方法
CN115259840B (zh) * 2022-08-31 2023-05-23 广东广青金属科技有限公司 一种耐火材料及其制备方法

Citations (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US4141722A (en) * 1976-02-10 1979-02-27 Osaka Iron & Steel Co., Ltd. Method of treating ferruginous slags
CN1435501A (zh) * 2002-12-13 2003-08-13 西安建筑科技大学 一种用镍熔炼炉渣和钢渣的混合渣炼铁的方法
CN104946840A (zh) * 2015-06-19 2015-09-30 西安建筑科技大学 利用含铁有色冶炼渣制备直接还原铁和胶凝性材料的方法
CN106755652A (zh) * 2016-12-10 2017-05-31 东北大学 一种含钛熔渣冶金一步法回收的方法
CN106755651A (zh) * 2016-12-10 2017-05-31 东北大学 一种含稀土和/或铌熔渣冶金一步法回收的方法

Family Cites Families (8)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN101100708A (zh) * 2006-07-07 2008-01-09 隋智通 从铜冶炼渣中分离铁与铜两种组分的方法
CN101886154B (zh) * 2010-07-02 2012-06-20 昆明理工大学 一种铜渣与铁矿石混合熔融还原制得低铜铁水的方法
AU2011279533B2 (en) * 2010-07-15 2014-08-07 Glencore Technology Pty Limited Pyrometallurgical method
CN104120209B (zh) * 2014-07-29 2016-08-24 钢铁研究总院 一种液态镍渣熔融还原生产含镍铁水的方法
CN104152714B (zh) * 2014-08-01 2016-03-30 西安建筑科技大学 一种硫化矿镍铜火法冶炼与提铁方法
CN105177295B (zh) * 2015-08-04 2018-01-02 江苏省冶金设计院有限公司 一种综合处理赤泥与铜渣的方法
CN105603213B (zh) * 2016-04-05 2017-08-25 江西同德盛元镍业有限公司 一种用镍废料生产水淬镍豆的方法
CN106755654A (zh) * 2016-12-10 2017-05-31 东北大学 一种熔渣冶金熔融还原生产的方法

Patent Citations (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
US4141722A (en) * 1976-02-10 1979-02-27 Osaka Iron & Steel Co., Ltd. Method of treating ferruginous slags
CN1435501A (zh) * 2002-12-13 2003-08-13 西安建筑科技大学 一种用镍熔炼炉渣和钢渣的混合渣炼铁的方法
CN104946840A (zh) * 2015-06-19 2015-09-30 西安建筑科技大学 利用含铁有色冶炼渣制备直接还原铁和胶凝性材料的方法
CN106755652A (zh) * 2016-12-10 2017-05-31 东北大学 一种含钛熔渣冶金一步法回收的方法
CN106755651A (zh) * 2016-12-10 2017-05-31 东北大学 一种含稀土和/或铌熔渣冶金一步法回收的方法

Non-Patent Citations (1)

* Cited by examiner, † Cited by third party
Title
ZHANG, JUHUA ET AL.: "Mechanism of Vanadium Slag Roasting with Calcium Oxide", INTERNATIONAL JOURNAL OF MINERAL PROCESSING, vol. 138, 18 March 2015 (2015-03-18), pages 20 - 29, XP055592257, ISSN: 0301-7516, DOI: 10.1016/j.minpro.2015.03.007 *

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