WO2020057026A1 - 一种从含溴冶炼烟灰中回收溴盐的方法 - Google Patents

一种从含溴冶炼烟灰中回收溴盐的方法 Download PDF

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WO2020057026A1
WO2020057026A1 PCT/CN2019/071756 CN2019071756W WO2020057026A1 WO 2020057026 A1 WO2020057026 A1 WO 2020057026A1 CN 2019071756 W CN2019071756 W CN 2019071756W WO 2020057026 A1 WO2020057026 A1 WO 2020057026A1
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lead
solution
leaching
zinc
bromine
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PCT/CN2019/071756
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French (fr)
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潘德安
吴玉锋
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北京工业大学
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Priority to US16/485,141 priority Critical patent/US10954125B2/en
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    • 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
    • C01—INORGANIC CHEMISTRY
    • C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
    • C01B9/00—General methods of preparing halides
    • C01B9/04—Bromides
    • C—CHEMISTRY; METALLURGY
    • C01—INORGANIC CHEMISTRY
    • C01D—COMPOUNDS OF ALKALI METALS, i.e. LITHIUM, SODIUM, POTASSIUM, RUBIDIUM, CAESIUM, OR FRANCIUM
    • C01D3/00—Halides of sodium, potassium or alkali metals in general
    • C01D3/10—Bromides
    • C—CHEMISTRY; METALLURGY
    • C01—INORGANIC CHEMISTRY
    • C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
    • C01G21/00—Compounds of lead
    • C01G21/20—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
    • C22B13/00—Obtaining lead
    • C22B13/04—Obtaining lead by wet processes
    • C22B13/045—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
    • 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
    • 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
    • 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
    • C22B7/008—Wet processes by an alkaline or ammoniacal leaching
    • 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
    • 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
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P20/00—Technologies relating to chemical industry
    • Y02P20/10—Process efficiency

Definitions

  • the invention relates to the field of high-efficiency separation and recovery of bromine by the full wet method, in particular to a method of high-efficiency separation of bromine salt and lead-zinc recovery by a two-step method of incineration of circuit board by soot.
  • Brominated flame retardants have good fire protection effects on plastics and textiles, and are one of the main chemical flame retardants. Therefore, a considerable part of waste electrical and electronic appliances are made of brominated flame retardant plastics. Whether such plastics can be recycled safely has become one of the environmental protection focuses of people's attention.
  • Broken flame retardant plastics currently available recycling methods are: 1 landfill; 2 mechanical recycling; 3 incineration without energy recovery; 4 incineration to recover energy; 5 incineration or cracking to recover bromine or hydrobromic acid.
  • This high-concentration alkali metal ionization medium solution can provide high chemical activity and high activity negative oxygen ions, which is called a sub-molten salt unconventional medium.
  • the medium has excellent physical and chemical properties such as low vapor pressure, high boiling point, and good fluidity, as well as excellent reaction separation characteristics such as high OH-activity coefficient, high chemical reaction activity, and adjustable separation function. Based on the above-mentioned superior characteristics, it can achieve a 100% leaching rate, and at the same time, zero emissions to the outside.
  • the purpose of the invention is mainly to solve the high-efficiency separation and recovery of bromine salts in bromine-containing smelting soot, and the efficient conversion and separation of lead and zinc, which has significant energy-saving and emission-reducing effects.
  • the method for recovering bromine salt from bromine-containing smelting soot according to the present invention is performed according to the following steps:
  • Sub-molten salt leaching The bromine-containing smelt soot and sodium hydroxide are sub-molten salt leached in a sodium hydroxide system, in which the solid-liquid ratio of the bromine-containing smelt soot: sodium hydroxide system is 1:20 to 1:40. Kg / l, the sodium hydroxide system is a sodium hydroxide solution with a mass concentration of 25 to 45%, the leaching temperature is 140 to 200 ° C, and the leaching time is 2 to 4 hours to obtain the leaching solution and leaching residue;
  • step (1) The leaching slag obtained in step (1) is washed, wherein the leaching slag: water-solid-liquid ratio is 1: 5 to 1:10 kg / liter, and water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag Centralized processing
  • Step (3) Lead-zinc separation: Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 1: 3 to 3: 1. Add 98% industrial concentrated sulfuric acid to the mass concentration until the pH of the combined solution reaches 6.5 ⁇ 8, and filter to obtain lead-separated zinc slag and lead-separated zinc solution;
  • step (3) The lead-containing zinc slag obtained in step (3) is mixed with water, and the solid-liquid ratio of the lead-containing zinc slag and water is 1: 1 to 1: 2 kg / liter. Stir and add a mass concentration of 98. % Industrial concentrated sulfuric acid, until the pH of the solution reaches 4.5 to 6, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the present invention adopts sodium hydroxide sub-molten salt leaching technology, and the reaction temperature is reduced by 400-500 ° C, which has good energy saving effect.
  • the present invention also adopts membrane separation and concentration technology, which will obtain The concentrated lead-zinc liquid is efficiently concentrated to reduce the energy consumption of subsequent bromine salt evaporation and crystallization, and the obtained clean water is returned to the cleaning process to avoid the generation of tail liquid and realize recycling.
  • Figure 1 shows a flow chart of a method for recovering bromine salts from bromine-containing smelting ashes
  • Sub-molten salt leaching bromine-containing smelted soot and sodium hydroxide are sub-molten salt leached in a sodium hydroxide system, in which the bromine-containing smelted soot: sodium hydroxide system has a solid-liquid ratio of 1:20 kg / liter, The sodium hydroxide system is a 25% sodium hydroxide solution, the leaching temperature is 140 ° C, and the leaching time is 2 hours to obtain an leaching solution and leaching residue;
  • step (2) Washing: washing the leaching slag obtained in step (1), wherein the leaching slag: water-solid-liquid ratio is 1: 5 kg / liter, and the water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 1: 3, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid, until the pH of the combined solution reaches 6.5, and filtered to obtain lead and zinc slag;
  • Lead separation The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 1 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the pH of the solution reaches 4.5, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 97.3%, the recovery rate of lead was 94.1%, and the recovery rate of zinc was 95.8%.
  • Leaching sub-molten salts of bromine-containing smelted soot and sodium hydroxide in a sodium hydroxide system in which the solid-liquid ratio of bromine-containing smelted soot: sodium hydroxide system is 1:40 kg / liter, The sodium hydroxide system is a 45% sodium hydroxide solution, the leaching temperature is 200 ° C, and the leaching time is 4 hours to obtain an leaching solution and leaching residue;
  • step (2) Washing: washing the leaching slag obtained in step (1), wherein the leaching slag: water-solid-liquid ratio is 1:10 kg / liter, and the water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 3: 1, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid until the combined solution reaches pH 8 and filtered to obtain lead-separated zinc slag and lead-separated zinc solution;
  • Lead separation The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 2 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the solution pH reaches 6, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 96.3%, the recovery rate of lead was 97.2%, and the recovery rate of zinc was 98.3%.
  • Sub-molten salt leaching bromine-containing smelt soot and sodium hydroxide are sub-molten salt leached in a sodium hydroxide system, in which the solid-liquid ratio of the bromine-containing smelt soot: sodium hydroxide system is 1:30 kg / liter,
  • the sodium hydroxide system is a sodium hydroxide solution with a concentration of 35% by mass, the leaching temperature is 180 ° C, and the leaching time is 3 hours to obtain an leaching solution and leaching residue;
  • step (1) The leaching slag obtained in step (1) is washed, wherein the leaching slag: water-solid-liquid ratio is 1: 7 kg / liter, and the water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leach solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leach solution to the cleaning solution is 1: 1, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid, until the pH of the combined solution reaches 7, and filtered to obtain lead-separated zinc slag and lead-separated zinc solution;
  • Lead separation The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 1.5 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the solution pH reaches 5, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 99.3%, the recovery rate of lead was 98.5%, and the recovery rate of zinc was 97.2%.
  • Sub-molten salt leaching bromine-containing smelted soot and sodium hydroxide are sub-molten salt leached in a sodium hydroxide system, in which the bromine-containing smelted soot: sodium hydroxide system has a solid-liquid ratio of 1:20 kg / liter, The sodium hydroxide system is a 45% sodium hydroxide solution, the leaching temperature is 140 ° C, and the leaching time is 4 hours to obtain an leaching solution and leaching residue;
  • step (2) Washing: washing the leaching slag obtained in step (1), wherein the leaching slag: water-solid-liquid ratio is 1: 5 kg / liter, and the water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 3: 1, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid, until the pH of the combined solution reaches 6.5, and filtered to obtain lead and zinc slag;
  • Lead separation The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 2 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the pH of the solution reaches 4.5, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 96.2%, the recovery rate of lead was 97.1%, and the recovery rate of zinc was 97.3%.
  • Leaching sub-molten salts of bromine-containing smelted soot and sodium hydroxide in a sodium hydroxide system in which the solid-liquid ratio of bromine-containing smelted soot: sodium hydroxide system is 1:40 kg / liter, The sodium hydroxide system is a 25% sodium hydroxide solution, the leaching temperature is 200 ° C, and the leaching time is 2 hours to obtain the leaching solution and leaching residue;
  • step (2) Washing: washing the leaching slag obtained in step (1), wherein the leaching slag: water-solid-liquid ratio is 1:10 kg / liter, and the water is at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 1: 3, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid until the combined solution reaches pH 8 and filtered to obtain lead-separated zinc slag and lead-separated zinc solution;
  • Lead separation The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 1 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the solution pH reaches 6, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 97.3%, the recovery rate of lead was 98.1%, and the recovery rate of zinc was 95.2%.
  • Sub-molten salt leaching bromine-containing smelt soot and sodium hydroxide are sub-molten salt leached in a sodium hydroxide system, in which the solid-liquid ratio of the bromine-containing smelt soot: sodium hydroxide system is 1:35 kg / liter,
  • the sodium hydroxide system is a 30% sodium hydroxide solution with a leaching temperature of 160 ° C and a leaching time of 2.5 hours to obtain an leaching solution and leaching residue;
  • step (2) Washing: washing the leaching slag obtained in step (1), wherein the leaching slag has a water-solid-liquid ratio of 1: 8 kg / liter and water at room temperature to obtain a washing liquid and a washing slag, and the washing slag is collectively processed;
  • Lead-zinc separation Combine the leaching solution obtained in step (1) with the cleaning solution obtained in step (2) to obtain a combined solution, wherein the volume ratio of the leaching solution to the cleaning solution is 1: 2, and add a mass concentration to the mixed solution. It is 98% industrial concentrated sulfuric acid, until the pH of the combined solution reaches 7.5, and filtered to obtain lead and zinc slag;
  • step (3) The lead-containing zinc slag obtained in step (3) is mixed with water.
  • the solid-liquid ratio of the lead-containing zinc slag and water is 1: 1.2 kg / liter. Stir and add 98% industrial concentrated sulfuric acid. Until the solution pH reaches 5.5, filtering to obtain lead sulfate and lead separation solution;
  • step (6) Evaporation and crystallization of zinc:
  • the lead separation liquid obtained in step (6) is subjected to evaporation and crystallization to obtain crude zinc sulfate.
  • the recovery rate of bromine salt was 96.3%, the recovery rate of lead was 95.8%, and the recovery rate of zinc was 98.1%.

Abstract

一种从含溴冶炼烟灰中回收溴盐的方法，属于烟灰全湿法回收溴盐及铅锌有价元素领域，特别涉及含溴冶炼烟灰采用亚熔盐法对溴盐高效分离回收和铅锌的转化分离。主要包括亚熔盐浸出、清洗、工业硫酸分步调pH分离提取铅和锌、膜分离浓缩回用清水、溴盐蒸发结晶回收粗溴盐等步骤。与传统烟灰焙烧回收工艺相比，采用了氢氧化钠亚熔盐浸出和膜分离浓缩技术，大大降低反应温度和尾液排放，具有很好节能减排的效果。

Description

一种从含溴冶炼烟灰中回收溴盐的方法 技术领域
本发明涉及溴的全湿法高效分离回收领域，特别涉及线路板焚烧烟灰两步法溴盐高效分离及铅锌回收的方法。
背景技术
回收利用废旧电子电气设备是人类面临的一大挑战。溴化阻燃剂在塑胶及纺织品上的防火效果很好，是主要的化学阻燃剂之一。因此，废旧电子电器有相当一部分是用溴系阻燃塑料制备的，这类塑料能否安全地回收利用已成了人们关注的环保焦点之一。溴系阻燃塑料目前可采用的回收方法有:1掩埋；2机械回收；3焚烧而不回收能量；4焚烧以回收能量；5焚烧或裂解以回收溴或氢溴酸。由溴系阻燃塑料回收溴或氢溴酸时，可采用几种工艺:1与城市固体垃圾共焚烧法；2裂解/气化法；3铜冶炼法。随着铜协同冶炼技术和装备的发展，铜冶炼法协同处置溴系阻燃塑料已经成为发展方向。含溴化阻燃剂的有机物进行焚烧或者协同冶炼，将产生大量含溴冶炼烟灰，其中主要成分是溴化钠。溴化钠是重要的溴盐，主要应用于感光胶片、医药(镇静剂)、农药、香料、染料等工业。同时，溴化钠溶于水、低毒，有刺激性，属于危险品。因此，冶炼烟灰中的含溴物质既有极高的资源型，同时具有一定的环境风险，必须合理回收。
由于含溴冶炼烟灰中含有溴化亚铜等物质，阻碍溴的分离，传统碱浸工艺无法有效回收溴。中科院过程工程研究所张懿等人，通过研究对比常规电解质溶液和熔盐介质之间在流体力学、热力学、化学动力学这些方面的共同点与差异点，提出了一种介于熔盐和常规电解质之间的高浓度电解质溶液，针对钒渣提钒的研究首次提出了亚熔盐回收技术。这种高浓度碱金属离子化介质溶液能够提供高化学活性和高活度负氧离子，被称作亚熔盐非常规介质。该介质具有蒸汽压低、沸点高、流动性好等优良物化性质和高OH-活度系数、高化学反应活性、分离功能可调等优良反应分离特性。基于以上所述优越特性，可以实现接100％浸出率，同时对外零排放污染。
发明内容
本发明的目的主要解决含溴冶炼烟灰中溴盐高效分离回收、铅锌高效转化分离，具有显著的节能减排效果。
本发明所述的一种从含溴冶炼烟灰中回收溴盐的方法按照如下步骤进行：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:20～1:40公斤/升，氢氧化钠体系为质量浓度为25～45％的氢氧化钠溶液，浸出温度为140～200℃，浸出时间为2～4小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:5～1:10公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:3～3:1，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到6.5～8，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1～1:2公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到4.5～6，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
与传统烟灰焙烧回收工艺相比，本发明采用了氢氧化钠亚熔盐浸出技术，反应温度降低400～500℃，具有很好节能效果，同时，本发明还采用了膜分离浓缩技术，将得到的分铅锌液进行高效浓缩，降低后续溴盐蒸发结晶能耗，得到的清水返回清洗工序，避免尾液产生，实现循环使用。
附图说明
图1表示一种从含溴冶炼烟灰中回收溴盐的方法流程图
具体实施方式
实施例1
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:20公斤/升，氢氧化钠体系为质量浓度为25％的氢氧化钠溶液，浸出温度为140℃，浸出时间为2小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:5公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:3，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到6.5，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到4.5，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率97.3％，铅回收率94.1％，锌回收率95.8％。
实施例2
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:40公斤/升，氢氧化钠体系为质量浓度为45％的氢氧化钠溶液，浸出温度为200℃，浸出时间为4小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:10公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为3:1，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到8，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:2公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到6，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率96.3％，铅回收率97.2％，锌回收率98.3％。
实施例3
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:30公斤/升，氢氧化钠体系为质量浓度为35％的氢氧化钠溶液，浸出温度为180℃，浸出时间为3小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:7公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:1，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到7，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进 行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1.5公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到5，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率99.3％，铅回收率98.5％，锌回收率97.2％。
实施例4
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:20公斤/升，氢氧化钠体系为质量浓度为45％的氢氧化钠溶液，浸出温度为140℃，浸出时间为4小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:5公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为3:1，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到6.5，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:2公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到4.5，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率96.2％，铅回收率97.1％，锌回收率97.3％。
实施例5
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:40公斤/升，氢氧化钠体系为质量浓度为25％的氢氧化钠溶液，浸出温度为200℃，浸出时间为2小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:10公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:3，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到8，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到6，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率97.3％，铅回收率98.1％，锌回收率95.2％。
实施例6
按照如下步骤进行处理：
(1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，其中含溴冶炼烟灰:氢氧化钠体系固液比为1:35公斤/升，氢氧化钠体系为质量浓度为30％的氢氧化钠溶液，浸出温度为160℃，浸出时间为2.5小时，得到浸出液和浸出渣；
(2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣: 水固液比为1:8公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
(3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:2，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到7.5，过滤得到分铅锌渣和分铅锌液；
(4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到浓缩液和清水，清水返回清洗工序；
(5)溴盐蒸发结晶：将步骤(4)得到的浓缩液进行蒸发结晶，得到粗溴盐；
(6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1.2公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到5.5，过滤得到硫酸铅和分铅液；
(7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
溴盐回收率96.3％，铅回收率95.8％，锌回收率98.1％。

Claims (2)

  1. 一种从含溴冶炼烟灰中回收溴盐的方法，其特征在于，具体步骤如下：
    (1)亚熔盐浸出：将含溴冶炼烟灰与氢氧化钠在氢氧化钠体系中进行亚熔盐浸出，得到浸出液和浸出渣；
    (2)清洗：将步骤(1)得到的浸出渣进行清洗，其中浸出渣:水固液比为1:5～1:10公斤/升，水为室温，得到清洗液和清洗渣，清洗渣集中处理；
    (3)分铅锌：将步骤(1)得到的浸出液与步骤(2)得到的清洗液合并，得到合并液，其中浸出液与清洗液的体积比为1:3～3:1，向混合液中加入质量浓度为98％工业浓硫酸，直到合并液pH到6.5～8，过滤得到分铅锌渣和分铅锌液；
    (4)膜分离浓缩：将步骤(3)得到的分铅锌液采用反渗透膜进行分离浓缩，得到脱铅锌液和清水，清水返回清洗工序；
    (5)溴盐蒸发结晶：将步骤(4)得到的脱铅锌液进行蒸发结晶，得到粗溴盐；
    (6)分铅：将步骤(3)得到的分铅锌渣与水进行混合，分铅锌渣与水的固液比1:1～1:2公斤/升，搅拌并加入质量浓度为98％工业浓硫酸，直到溶液pH到4.5～6，过滤得到硫酸铅和分铅液；
    (7)锌蒸发结晶：将步骤(6)得到的分铅液进行蒸发结晶，得到粗硫酸锌。
  2. 如权利要求1所述的一种从含溴冶炼烟灰中回收溴盐的方法，其特征在于，步骤(1)中含溴冶炼烟灰:氢氧化钠体系固液比为1:20～1:40公斤/升，氢氧化钠体系为质量浓度为25～45％的氢氧化钠溶液，浸出温度为140～200℃，浸出时间为2～4小时。
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