WO2021129205A1 - 一种废线路板裂解渣与冶炼烟灰协同处置的方法 - Google Patents

一种废线路板裂解渣与冶炼烟灰协同处置的方法 Download PDF

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WO2021129205A1
WO2021129205A1 PCT/CN2020/127973 CN2020127973W WO2021129205A1 WO 2021129205 A1 WO2021129205 A1 WO 2021129205A1 CN 2020127973 W CN2020127973 W CN 2020127973W WO 2021129205 A1 WO2021129205 A1 WO 2021129205A1
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copper
circuit board
leaching
roasting
zinc
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French (fr)
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吴玉锋
刘功起
李彬
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Beijing University of Technology
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Beijing University of Technology
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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
    • 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
    • C22B7/02—Working-up flue dust
    • C—CHEMISTRY; METALLURGY
    • C01—INORGANIC CHEMISTRY
    • C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
    • C01G9/00—Compounds of zinc
    • C01G9/06—Sulfates
    • 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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/06—Sulfating roasting
    • 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
    • C22B11/04—Obtaining noble metals by wet processes
    • C22B11/042—Recovery of noble metals 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
    • C22B11/00—Obtaining noble metals
    • C22B11/04—Obtaining noble metals by wet processes
    • C22B11/042—Recovery of noble metals from waste materials
    • C22B11/044—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
    • C22B15/00—Obtaining copper
    • C22B15/0063—Hydrometallurgy
    • C22B15/0065—Leaching or slurrying
    • C22B15/0067—Leaching or slurrying with acids or salts thereof
    • C22B15/0071—Leaching or slurrying with acids or salts thereof containing sulfur
    • 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
    • C22B15/0063—Hydrometallurgy
    • C22B15/0084—Treating solutions
    • C22B15/0089—Treating solutions by chemical methods
    • 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
    • C22B19/04—Obtaining zinc by distilling
    • 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
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • 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
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
    • C22B3/46—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes by substitution, e.g. by cementation
    • 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
    • C22B7/005—Separation by a physical processing technique only, e.g. by mechanical breaking
    • 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
    • C22B7/006—Wet processes
    • C22B7/007—Wet processes by acid leaching
    • 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
    • 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
    • Y02W—CLIMATE CHANGE MITIGATION TECHNOLOGIES RELATED TO WASTEWATER TREATMENT OR WASTE MANAGEMENT
    • Y02W30/00—Technologies for solid waste management
    • Y02W30/50—Reuse, recycling or recovery technologies
    • Y02W30/82—Recycling of waste of electrical or electronic equipment [WEEE]

Definitions

  • the invention relates to the field of co-processing typical ash and slag of waste circuit boards by fire method, and particularly relates to a method for coupling and synergistic debromination of waste circuit board cracking slag and waste circuit board smelting soot and comprehensive recovery of copper and zinc.
  • Waste circuit boards are components with high recycling value produced in the process of dismantling waste electronic appliances, and their disposal is the core of the high-value utilization of electronic appliances.
  • the flame retardant contained in waste circuit boards is generally brominated epoxy resin, tetrabromobisphenol A reactive bromine-containing flame retardant, with a bromine content of 5-15%.
  • Such flame retardants and other molecules are generally chemically bonded to the circuit board
  • the resin materials in the circuit board are combined together, so under the normal use conditions of the circuit board, the brominated flame retardant is safe and stable, but in the recycling process of the circuit board, if it is not handled properly, it is easy to cause the loss of the brominated flame retardant, Decomposes and causes serious environmental pollution.
  • bromine is an important chemical raw material. Therefore, how to recover bromine from bromine-containing flame retardants is a problem that needs to be solved in the current waste circuit board processing industry and the entire flame retardant electronic plastic recycling field. .
  • Most of the resins in waste circuit boards are thermosetting materials, and general physical melting and regeneration techniques cannot be used for recycling.
  • Fire technologies such as incineration, smelting, and pyrolysis are considered to be the most effective technology for recycling waste circuit boards.
  • the fire disposal process often produces certain solid residues, such as waste circuit board cracking slag and smelting/incineration soot.
  • the traditional recovery methods of copper-containing smelting soot mainly include fire method and wet method.
  • the fire method is mainly used in the smelting furnace.
  • the volatile lead, zinc, arsenic, etc. are volatilized and enriched in a reducing atmosphere. Others Elements enter into polymetallic residues.
  • the comprehensive recovery rate of valuable metals in this way is low. Improper treatment of the generated flue gas seriously affects the environment and the health of workers.
  • wet recovery is typically acid leaching -Chemical transformation and other methods, mainly for copper, zinc, nickel and other base metals that are easily soluble in acids and alkalis.
  • the obtained leachate is enriched by extraction and stripping or displacement precipitation.
  • the outstanding disadvantage of this method is the stock of tail liquid. Large and difficult to recycle.
  • the typical recovery method for waste circuit board cracking slag is often through mechanical pulverization and pretreatment, and then use magnetic separation, eddy current separation, or electrical separation to completely separate the metal and non-metal components, and then the recovered metal can be recovered by fire method or Wet further treatment, although this method can be the initial separation of metals and non-metals in the pyrolysis slag, it has the problems of incomplete separation and low comprehensive utilization rate.
  • the present invention proposes a process route of crushing and sorting, mixed roasting, intensified leaching, replacement of silver immersion, copper sulfide, and evaporation crystallization to co-process waste circuit board cracking slag and smelting soot.
  • the process treatment method It is simple and easy to implement, realizes the high-efficiency removal of bromide and the high-value utilization of copper and zinc. At the same time, there is no tail liquid discharge in the recovery process, which has significant environmental benefits.
  • the purpose of the present invention is mainly to solve the problem of the co-processing of waste circuit board cracking slag and smelting soot, and to propose a method for co-processing of cracking slag and smelting soot, debromination, enhanced copper and zinc leaching and high-value utilization.
  • the processing method has a short process flow and simple equipment, realizes high-efficiency removal of bromide and high-value utilization of copper and zinc, and at the same time, there is no tail liquid discharge in the recovery process, which has significant environmental and economic benefits.
  • (1) Crushing and sorting the waste circuit board cracked slag is crushed to particles with a particle size of 0.5 to 4.5 mm with a shear crusher, and the metal and non-metals are separated by electrostatic or shaker sorting to obtain glass Fibers and sorting residues, glass fibers are centrally disposed of, and the content of glass fibers in the sorting residues is 0-5%;
  • step (3) Intensified leaching: The calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch-type ultrasonic reactor, where the sulfuric acid concentration is 1-20g/L, or tap water is directly used as the leaching agent.
  • the mass ratio is 3:1 ⁇ 6:1, the leaching temperature is 25 ⁇ 65°C, the leaching time is 1.0 ⁇ 2.5 hours, the stirring speed is 180 ⁇ 360rpm, and the ratio of ultrasonic working time to intermittent time is (10 ⁇ 20):10( min/min), the ultrasonic power is 25 ⁇ 55Wh per kilogram of slurry, the frequency is 5 ⁇ 30KHz, the leaching residue and the copper-zinc leaching solution are obtained, and the leaching residue is returned to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 0.3-1.5g of copper powder is added to each liter of copper-zinc leaching solution, the reaction temperature is 50-80°C, and the reaction time is 1.0 ⁇ 3.0 hours to obtain crude silver and immersion silver post-liquid;
  • Copper sulfide sinking pass the post-silver sinking liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 0.5 to 2.0 L/min, and hydrogen sulfide is fed into the copper sinking liquid per liter of hydrogen sulfide. ⁇ 20L, to obtain crude copper sulfide and copper immersion liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the present invention adopts the co-processing of waste circuit board cracking slag and smelting soot, it solves the problem of the large difference in bromine content in the typical ash slag of the two pyrotechnical treatment of waste circuit boards.
  • the common coupling it achieves a The purpose of the process to dispose of two solid wastes.
  • the sulphated roasting not only meets the requirement of high-efficiency removal of waste circuit board cracking slag and smelting soot bromide, but also realizes the purpose of selective conversion of base metal copper and zinc. Through enhanced leaching and selective precipitation, high-value products are obtained.
  • the co-processing of these two kinds of wastes provides a reference value for the recycling of other low-grade secondary resources.
  • the invention is particularly suitable for the treatment of waste circuit board cracking slag and smelting soot with a large difference in inorganic bromide content.
  • Roasting and synergistic debromination, enhanced leaching of base metals, directional enrichment and high-value utilization are featured. It has short process flow, simple equipment, and no The discharge of tail liquid is characterized by significant economic and environmental benefits.
  • Figure 1 shows a process flow diagram of a method for co-processing waste circuit board cracking slag and smelting soot
  • step (2) Roasting of mixed material:
  • the separation residue obtained in step (1) is mixed with waste circuit board smelting soot to obtain a mixed material.
  • the mixed material contains 16.2% bromine, 32.1% copper, and 18.6% of zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, and tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 0.5 kg of waste circuit board smelting soot, and the mass ratio of the mixture to the concentrated sulfuric acid is 2:1.
  • No tap water is added, the roasting temperature is 250°C, and the roasting time is 1.5 hours to obtain roasting sand and roasting flue gas.
  • the roasting flue gas is absorbed by lye and then returned to the sodium bromide purification process;
  • step (3) Intensified leaching:
  • the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor, and tap water is used as the leaching agent.
  • the liquid-to-solid mass ratio is 3:1 and the leaching temperature is 25°C, the leaching time is 1.0 hour, the stirring speed is 180rpm, the ratio of ultrasonic working time to intermittent time is 10:10 (min/min), the ultrasonic power is 25Wh per kilogram of slurry, the frequency is 5KHz, and the leaching residue and copper zinc are obtained.
  • the leaching solution and leaching residue are returned to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 0.3g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 50°C, the reaction time is 1.0 hour, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide sinking pass the post-silver sinking liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 0.5L/min, and 5L hydrogen sulfide per liter of copper sinking liquid is passed into it to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 99.1%, the recovery rate of copper was 98.5%, and the recovery rate of zinc was 97.0%.
  • roasting of the mixture the separation residue obtained in step (1) is mixed with the waste circuit board smelting soot to obtain a mixture.
  • the mixture contains 19.4% bromine, 35.5% copper, and 17.0% zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, while tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 3.0 kg of waste circuit board smelting soot, and the mass ratio of the mixture to the concentrated sulfuric acid is 1:2.
  • the amount of tap water added is 20% of the mass of the mixture, the roasting temperature is 450°C, and the roasting time is 3.5 hours to obtain roasting sand and roasting flue gas.
  • the roasting flue gas is absorbed by lye and returned to the sodium bromide purification process;
  • step (3) Intensified leaching:
  • the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor.
  • the sulfuric acid concentration is 20g/L
  • the liquid-to-solid mass ratio is 6:1
  • the leaching temperature is 65°C
  • the leaching time is 2.5 hours
  • the stirring speed is 360rpm
  • the ratio of ultrasonic working time to intermittent time is 20:10 (min/min)
  • the ultrasonic power is 55Wh per kilogram of slurry
  • the frequency is 30KHz
  • the leaching residue and copper and zinc are obtained.
  • the leaching solution and leaching residue are returned to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 1.5g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 80°C, the reaction time is 3.0 hours, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide precipitation pass the post-precipitation liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide aeration flow rate is 2.0L/min, and 20L hydrogen sulfide per liter of post-precipitation copper liquid is passed to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 99.6%, the recovery rate of copper was 98.4%, and the recovery rate of zinc was 97.9%.
  • step (2) Roasting of mixed material:
  • the separation residue obtained in step (1) is mixed with waste circuit board smelting soot to obtain a mixed material.
  • the mixed material contains 22.9% bromine, 36.8% copper, and 19.1% zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, while tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 1.0 kg of waste circuit board smelting soot, and the mass ratio of the mixture to concentrated sulfuric acid is 1:1.
  • the amount of tap water added is 1% of the mass of the mixture, the roasting temperature is 300°C, and the roasting time is 2.0 hours to obtain roasted sand and roasting flue gas.
  • the roasting flue gas is absorbed with lye and returned to the sodium bromide purification process;
  • step (3) Intensified leaching:
  • the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor.
  • the sulfuric acid concentration is 1g/L
  • the liquid-to-solid mass ratio is 4:1
  • the leaching temperature is 30°C
  • the leaching time is 1.5 hours
  • the stirring speed is 200rpm
  • the ratio of ultrasonic working time to intermittent time is 12:10 (min/min)
  • the ultrasonic power is 30Wh per kilogram of slurry
  • the frequency is 10KHz
  • the leaching residue and copper and zinc are obtained.
  • the leaching solution and leaching residue return to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 0.5g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 55°C, the reaction time is 1.5 hours, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide precipitation pass the post-precipitation liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 0.8L/min, and 8L of hydrogen sulfide per liter of post-precipitation copper solution is passed to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 98.6%
  • the recovery rate of copper was 97.3%
  • the recovery rate of zinc was 97.7%.
  • roasting of the mixture the separation residue obtained in step (1) is mixed with the waste circuit board smelting soot to obtain a mixture.
  • the mixture contains 24.7% of bromine, 38.3% of copper, and 19.7% of zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, while tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 2.5 kg of waste circuit board smelting soot, and the mass ratio of the mixture to concentrated sulfuric acid is 1:2.
  • the amount of tap water added is 15% of the mass of the mixture, the roasting temperature is 400°C, and the roasting time is 3.0 hours to obtain roasting sand and roasting flue gas.
  • the roasting flue gas is absorbed by lye and returned to the sodium bromide purification process;
  • step (3) Intensified leaching:
  • the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor.
  • the sulfuric acid concentration is 15g/L
  • the liquid-to-solid mass ratio is 5:1
  • the leaching temperature is 60°C
  • leaching time is 2.0 hours
  • stirring speed is 340rpm
  • the ratio of ultrasonic working time to intermittent time is 18:10 (min/min)
  • ultrasonic power is 50Wh per kilogram of slurry
  • frequency is 25KHz
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 1.2g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 75°C, the reaction time is 2.5 hours, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide sinking pass the silver sinking liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 1.8L/min, and 18L hydrogen sulfide per liter of copper sinking liquid is passed into it to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 99.4%, the recovery rate of copper was 98.0%, and the recovery rate of zinc was 97.3%.
  • roasting of the mixture the separation residue obtained in step (1) is mixed with the waste circuit board smelting soot to obtain a mixture.
  • the mixture contains 18.3% bromine, 25.9% copper, and 18.2% zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, while tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 1.5 kg of waste circuit board smelting soot, and the mass ratio of the mixture to the concentrated sulfuric acid is 2:1.
  • the amount of tap water added is 10% of the mass of the mixture, the roasting temperature is 320°C, and the roasting time is 2.5 hours to obtain roasting sand and roasting flue gas.
  • the roasting flue gas is absorbed by lye and returned to the sodium bromide purification process;
  • step (3) Intensified leaching:
  • the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor.
  • the sulfuric acid concentration is 5g/L
  • the liquid-to-solid mass ratio is 3:1
  • the leaching temperature is 35°C
  • the leaching time is 1.5 hours
  • the stirring speed is 260rpm
  • the ratio of ultrasonic working time to intermittent time is 14:10 (min/min)
  • the ultrasonic power is 35Wh per kilogram of slurry
  • the frequency is 15KHz
  • the leaching residue and copper and zinc are obtained.
  • the leaching solution and leaching residue return to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 0.8g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 60°C, the reaction time is 2.5 hours, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide precipitation pass the post-silver precipitation liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 1.0L/min, and 10L hydrogen sulfide per liter of copper-precipitated liquid is passed into it to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 98.6%
  • the recovery rate of copper was 98.8%
  • the recovery rate of zinc was 98.1%.
  • roasting of the mixture the separation residue obtained in step (1) is mixed with the waste circuit board smelting soot to obtain a mixture.
  • the mixture contains 22.2% bromine, 31.8% copper, and 19.3% zinc.
  • Concentrated sulfuric acid with a mass fraction of 98% is added to the material, and tap water is added for dilution, and then sulfuric acid is roasted.
  • Each kilogram of the sorting residue is mixed with 2.0 kg of waste circuit board smelting soot, and the mass ratio of the mixture to the concentrated sulfuric acid is 1:1.
  • the amount of tap water added is 5% of the mass of the mixture, the roasting temperature is 350°C, and the roasting time is 3.5 hours to obtain roasting sand and roasting flue gas.
  • the roasting flue gas is absorbed by lye and returned to the sodium bromide purification process;
  • step (3) Intensified leaching: the calcined sand and dilute sulfuric acid solution obtained in step (2) are mechanically stirred and leached in a batch ultrasonic reactor, where the sulfuric acid concentration is 10g/L, the liquid-to-solid mass ratio is 5:1, and the leaching temperature is 55°C, leaching time is 2.0 hours, stirring speed is 300rpm, the ratio of ultrasonic working time to intermittent time is 15:10 (min/min), ultrasonic power is 40Wh per kilogram of slurry, frequency is 20KHz, and leaching residue and copper and zinc are obtained.
  • the leaching solution and leaching residue are returned to the process of enriching precious metals;
  • step (3) the copper-zinc leaching solution obtained in step (3) is added to copper powder for heating and stirring, wherein 1.0g of copper powder is added per liter of copper-zinc leaching solution, the reaction temperature is 70°C, the reaction time is 2.0 hours, and the crude Silver and immersion silver post-liquid;
  • Copper sulfide precipitation pass the post-precipitation liquid obtained in step (4) into excess hydrogen sulfide gas, where the hydrogen sulfide gas flow rate is 1.5L/min, and 15L hydrogen sulfide per liter of post-precipitation copper liquid is passed to obtain Crude copper sulfide and copper sinking liquid;
  • Zinc extraction by crystallization the copper precipitation solution obtained in step (5) is evaporated and crystallized to obtain crude zinc sulfate and a crystallization mother liquor, and the crystallization mother liquor is returned to strengthen the leaching process.
  • the removal rate of bromine was 99.0%, the recovery rate of copper was 99.1%, and the recovery rate of zinc was 98.7%.

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Abstract

一种废线路板裂解渣与冶炼烟灰协同处置的方法，属于火法处置废线路板典型灰渣有价元素综合回收利用领域，特别涉及废线路板裂解渣与废线路板冶炼烟灰协同处置脱溴与综合回收铜、锌的方法。其主要步骤如下：破碎分选、混料焙烧、强化浸出、置换沉银、硫化沉铜及蒸发结晶。与传统回收技术相比，该方法实现了一种回收工艺耦合处置两种固体废弃物的目的。通过混合硫酸焙烧既达到了废线路板裂解渣与冶炼烟灰溴化物协同脱除的要求，又实现了贱金属铜锌选择性转化的目的。通过强化浸出及选择性沉淀，得到了高值化产品。其具有工艺流程短，设备简单，无尾液排放，经济和环境效益显著的特点。

Description

一种废线路板裂解渣与冶炼烟灰协同处置的方法 技术领域
本发明涉及火法处置废线路板典型灰渣协同处置的领域，特别涉及废线路板裂解渣与废线路板冶炼烟灰耦合协同脱溴与综合回收铜、锌的方法。
背景技术
废线路板是废电子电器拆解过程中产生的具有较高回收价值的部件，其处置是电子电器高值化利用的核心。废线路板含有阻燃剂一般是溴化环氧树脂、四溴双酚A类反应型含溴阻燃剂，溴含量5～15％，这类阻燃剂等分子一般是通过化学键与线路板中的树脂材料结合在一起，因而在线路板正常使用条件下，溴化阻燃剂是安全和稳定的，但在线路板回收过程中，如果处理不当很容易造成溴化阻燃剂的流失、分解并导致严重的环境污染，同时溴是一种重要的化工原料，因此如何从含溴阻燃剂中回收这些溴是目前废旧线路板处理行业乃至整个阻燃剂电子塑料回收领域需要解决的问题。废线路板中的树脂多为热固性材料，不能采用一般的物理熔融再生技术回收处理，焚烧、冶炼、热解等火法技术被认为是循环利用废线路板最有效的技术。但是火法处置过程往往产生一定的固体残余物，典型的如废线路板裂解渣及冶炼/焚烧烟灰。这些典型灰渣不仅含有大量的铜、铅、锌、金、银等有价金属，同时受加热条件的影响，往往含有难处理的无机溴化物，主要成分如表1所示，因此废线路板冶裂解渣和冶炼烟灰具有一定的共性问题。但其中的这些金属溴化物化学性质稳定，难以采用传统的酸、碱、盐进行分离回收，因此，在回收废线路板裂解渣及冶炼烟灰中有价金属的同时，有必要预先脱除其中溴化物。
表1废线路板裂解渣和冶炼烟灰元素含量
Figure PCTCN2020127973-appb-000001
Figure PCTCN2020127973-appb-000002
含铜冶炼烟灰传统的回收方法主要有火法和湿法回收两种，火法回收主要是在熔炼炉中，将易挥发性的铅、锌、砷等在还原性气氛中挥发富集，其他元素进入多金属残余物，这种方式有价金属综合回收率较低，产生的烟气处理不当严重影响环境及劳动者身体健康，同时存在二次烟灰处置问题；湿法回收典型的是酸浸-化学转化等方法，主要是针对铜、锌、镍等易于溶于酸碱的贱金属，对得到的浸出液采用萃取反萃或置换沉淀等方式得到富集，这种方法突出缺点是尾液存量大且难以循环利用。
对于废线路板裂解渣典型的回收方法往往是通过机械粉碎预处理后，采用磁选、涡流分选、或电选的方式彻底分离金属和非金属成分，然后对于回收的金属可通过火法或湿法进一步处理，这种方式虽然能够是裂解渣中的金属和非金属初步分离，但存在分离不彻底，且综合利用率低的问题。
综合以上回收方法的特点，现有的单纯回收有价金属的冶金方法不能解决废线路板裂解渣及冶炼烟灰存在难处理无机金属溴化物的共性问题，因此有必要开发新工艺解决废线路板冶裂解渣和冶炼烟灰协同处置的问题。为了解决这一问题，本发明提出破碎分选、混料焙烧、强化浸出、置换沉银、硫化沉铜及蒸发结晶的工艺路线用以协同处置废线路板裂解渣及冶炼烟灰，该工艺处理方法简单易行，实现了溴化物高效脱除以及铜和锌的高值化利用，同时回收过程无尾液排放，具有显著的环境效益。
发明内容
本发明的目的主要解决废线路板裂解渣与冶炼烟灰协同处置的问题，提出一种裂解渣与冶炼烟灰协同焙烧脱溴、铜锌强化浸出及高值化利用的方法。该处理方法工艺流程较短，设备简单，实现了溴化物高效脱除以及铜和锌的高值化利用，同时回收过程无尾液排放，具有显著的环境和经济效益。
本发明所述的一种废线路板裂解渣与冶炼烟灰协同处置的方法按如下步骤进行：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为0.5～4.5mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量0～5％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到 混合料，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰0.5～3.0公斤，混合料与硫酸质量比为2:1～1:2，自来水添加量为混合料质量的1～20％，或不添加自来水，焙烧温度为250～450℃，焙烧时间为1.5～3.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为1～20g/L，或直接用自来水做浸出剂，液固质量比为3:1～6:1，浸出温度为25～65℃，浸出时间为1.0～2.5小时，搅拌速度为180～360rpm，超声工作时间与间歇时间比例为(10～20):10(min/min)，超声功率为每公斤料浆25～55Wh，频率为5～30KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉0.3～1.5g，反应温度为50～80℃，反应时间为1.0～3.0小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为0.5～2.0L/min，每升沉铜后液通入硫化氢5～20L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
与现有技术相比，由于本发明采用废线路板裂解渣与冶炼烟灰协同处置，解决了两种火法处置废线路板典型灰渣溴含量差距大的问题，通过共性耦合，达到了一种工艺处置两种固体废弃物的目的。且硫酸化焙烧既达到了废线路板裂解渣与冶炼烟灰溴化物高效脱除的要求，又实现了贱金属铜锌选择性转化的目的。通过强化浸出及选择性沉淀，得到了高值化产品。这两种废弃物的协同处置，为其他低品位二次资源的回收提供了借鉴参考价值。
本发明特别适合处置无机溴化物含量差距较大的废线路板裂解渣及冶炼烟灰的焙烧协同脱溴、贱金属强化浸出及定向富集及高值化利用，具有工艺流程短，设备简单，无尾液排放，经济和环境效益显著的特点。
附图说明
图1表示一种废线路板裂解渣与冶炼烟灰协同处置的方法工艺流程图
具体实施方式
以下结合实例旨在进一步说明本发明，而非限制本发明。
实施例1
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为0.5mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量为0％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴16.2％、铜32.1％、锌18.6％，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰0.5公斤，混合料与浓硫酸质量比为2:1，不添加自来水，焙烧温度为250℃，焙烧时间为1.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，直接用自来水做浸出剂，其中液固质量比为3:1，浸出温度为25℃，浸出时间为1.0小时，搅拌速度为180rpm，超声工作时间与间歇时间比例为10:10(min/min)，超声功率为每公斤料浆25Wh，频率为5KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉0.3g，反应温度为50℃，反应时间为1.0小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为0.5L/min，每升沉铜后液通入硫化氢5L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为99.1％，铜的回收率为98.5％，锌的回收率为97.0％。
实施例2
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为4.5mm的颗粒，采 用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量降低至5％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴19.4％、铜35.5％、锌17.0％，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰3.0公斤，混合料与浓硫酸质量比为1:2，自来水添加量为混合料质量的20％，焙烧温度为450℃，焙烧时间为3.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为20g/L，液固质量比为6:1，浸出温度为65℃，浸出时间为2.5小时，搅拌速度为360rpm，超声工作时间与间歇时间比例为20:10(min/min)，超声功率为每公斤料浆55Wh，频率为30KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉1.5g，反应温度为80℃，反应时间为3.0小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为2.0L/min，每升沉铜后液通入硫化氢20L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为99.6％，铜的回收率为98.4％，锌的回收率为97.9％。
实施例3
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为1.0mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量降低至1.5％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴22.9％、铜36.8％、锌19.1％，得到的混合料加入质量分数为98％ 的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰1.0公斤，混合料与浓硫酸质量比为1:1，自来水添加量为混合料质量的1％，焙烧温度为300℃，焙烧时间为2.0小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为1g/L，液固质量比为4:1，浸出温度为30℃，浸出时间为1.5小时，搅拌速度为200rpm，超声工作时间与间歇时间比例为12:10(min/min)，超声功率为每公斤料浆30Wh，频率为10KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉0.5g，反应温度为55℃，反应时间为1.5小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为0.8L/min，每升沉铜后液通入硫化氢8L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为98.6％，铜的回收率为97.3％，锌的回收率为97.7％。
实施例4
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为4.0mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量降低至4.5％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴24.7％、铜38.3％、锌19.7％，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰2.5公斤，混合料与浓硫酸质量比为1:2，自来水添加量为混合料质量的15％，焙烧温度为400℃，焙烧时间为3.0小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为15g/L，液固质量比为5:1，浸出温度为60℃，浸出时间为2.0小时，搅拌速度为340rpm，超声工作时间与间歇时间比例为18:10(min/min)，超声功率为每公斤料浆50Wh，频率为25KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉1.2g，反应温度为75℃，反应时间为2.5小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为1.8L/min，每升沉铜后液通入硫化氢18L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为99.4％，铜的回收率为98.0％，锌的回收率为97.3％。
实施例5
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为2.0mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量降低至2.5％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴18.3％、铜25.9％、锌18.2％，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰1.5公斤，混合料与浓硫酸质量比为2:1，自来水添加量为混合料质量的10％，焙烧温度为320℃，焙烧时间为2.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为5g/L，液固质量比为3:1，浸出温度为35℃，浸出时间为1.5小时，搅拌速度为260rpm，超声工作时间与间歇时间比例为14:10(min/min)，超声功率为每公斤料浆35Wh，频率为15KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工 序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉0.8g，反应温度为60℃，反应时间为2.5小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通气流量为1.0L/min，每升沉铜后液通入硫化氢10L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为98.6％，铜的回收率为98.8％，锌的回收率为98.1％。
实施例6
按照如下步骤进行回收：
(1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为3.0mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置，其中分选残余物中的玻璃纤维含量降低至3.0％；
(2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，混合料中含溴22.2％、铜31.8％、锌19.3％，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，其中每公斤分选残余物混入废线路板冶炼烟灰2.0公斤，混合料与浓硫酸质量比为1:1，自来水添加量为混合料质量的5％，焙烧温度为350℃，焙烧时间为3.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
(3)强化浸出：将步骤(2)得到的焙烧砂与稀硫酸溶液在间歇式超声反应器中机械搅拌浸出，其中硫酸浓度为10g/L，液固质量比为5:1，浸出温度为55℃，浸出时间为2.0小时，搅拌速度为300rpm，超声工作时间与间歇时间比例为15:10(min/min)，超声功率为每公斤料浆40Wh，频率为20KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
(4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌，其中每升铜锌浸出液加入铜粉1.0g，反应温度为70℃，反应时间为2.0小时，得到粗银和沉银后液；
(5)硫化沉铜：将步骤(4)得到的沉银后液通入过量的硫化氢气体，其中硫化氢通 气流量为1.5L/min，每升沉铜后液通入硫化氢15L，得到粗硫化铜和沉铜后液；
(6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
溴的脱除率为99.0％，铜的回收率为99.1％，锌的回收率为98.7％。
以上实施例仅用于说明本发明的优选实施方式，但本发明并不限于上述实施方式，在所述领域技术人员所具备的知识范围，在不违背科学及本发明思想情况下，在本发明的精神和原则之内所作的修改、等同替代及改进等，均应视为本申请的保护范围。

Claims (5)

  1. 一种废线路板裂解渣与冶炼烟灰协同处置的方法，其特征在于，具体步骤如下：
    (1)破碎分选：将废线路板裂解渣用剪切式破碎机破碎至粒径为0.5～4.5mm的颗粒，采用静电或摇床分选的方式使金属和非金属得到分离，得到玻璃纤维和分选残余物，玻璃纤维集中处置；
    (2)混料焙烧：将步骤(1)得到的分选残余物与废线路板冶炼烟灰进行混合后得到混合料，得到的混合料加入质量分数为98％的浓硫酸，同时加入自来水进行稀释，然后进行硫酸焙烧，焙烧温度为250～450℃，焙烧时间为1.5～3.5小时，得到焙烧砂和焙烧烟气，焙烧烟气用碱液吸收后返溴化钠提纯工序；
    (3)强化浸出：将步骤(2)得到的焙烧砂与硫酸溶液在间歇式超声反应器中机械搅拌浸出，液固质量比为3:1～6:1，浸出温度为25～65℃，浸出时间为1.0～2.5小时，超声功率为每公斤料浆25～55Wh，频率为5～30KHz，得到浸出渣和铜锌浸出液，浸出渣返富集贵金属工序；
    (4)置换沉银：将步骤(3)得到的铜锌浸出液加入铜粉进行加热搅拌置换反应，每升铜锌浸出液加入铜粉0.3～1.5g，反应温度为50～80℃，反应时间为1.0～3.0小时，得到粗银和沉银后液；
    (5)硫化沉铜：将步骤(4)得到的沉银后液通入硫化氢气体，得到粗硫化铜和沉铜后液；
    (6)结晶提锌：将步骤(5)得到沉铜后液进行蒸发结晶，得到粗硫酸锌和结晶母液，结晶母液返强化浸出工序。
  2. 如权利要求1所述的一种废线路板裂解渣与冶炼烟灰协同处置的方法，其特征在于，步骤(1)中分选残余物中的玻璃纤维含量为0～5％。
  3. 如权利要求1所述的一种废线路板裂解渣与冶炼烟灰协同处置的方法，其特征在于，步骤(2)中每公斤分选残余物混入废线路板冶炼烟灰0.5～3.0公斤，混合料与硫酸质量比为2:1～1:2，自来水添加量为混合料质量的1～20％，或不添加自来水。
  4. 如权利要求1所述的一种废线路板裂解渣与冶炼烟灰协同处置的方法，其特征在于，步骤(3)中硫酸浓度为1～20g/L，或直接用自来水做浸出剂，搅拌速度为180～360rpm，超声工作时间与间歇时间比例为(10～20):10。
  5. 如权利要求1所述的一种废线路板裂解渣与冶炼烟灰协同处置的方法，其特征在于，步骤(5)中硫化氢通气流量为0.5～2.0L/min，每升沉铜后液通入硫化氢5～20L。
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