WO2012083583A1 - 一种从白钨矿中提取钨的方法 - Google Patents

一种从白钨矿中提取钨的方法 Download PDF

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WO2012083583A1
WO2012083583A1 PCT/CN2011/001895 CN2011001895W WO2012083583A1 WO 2012083583 A1 WO2012083583 A1 WO 2012083583A1 CN 2011001895 W CN2011001895 W CN 2011001895W WO 2012083583 A1 WO2012083583 A1 WO 2012083583A1
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acid
scheelite
phosphoric acid
leaching
tungsten
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English (en)
French (fr)
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赵中伟
李江涛
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Central South University
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Central South University
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Priority claimed from CN2010106051102A external-priority patent/CN102021329B/zh
Priority claimed from CN2010106051070A external-priority patent/CN102021328B/zh
Priority claimed from CN2010106050951A external-priority patent/CN102080157B/zh
Priority claimed from CN2010106050947A external-priority patent/CN102021327B/zh
Priority claimed from CN2010106051032A external-priority patent/CN102080161B/zh
Priority to DE112011104540.9T priority Critical patent/DE112011104540B4/de
Priority to CA2815708A priority patent/CA2815708C/en
Priority to RU2013125145/02A priority patent/RU2532767C1/ru
Application filed by Central South University filed Critical Central South University
Publication of WO2012083583A1 publication Critical patent/WO2012083583A1/zh
Priority to US13/794,844 priority patent/US8771617B2/en
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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
    • C22B34/00—Obtaining refractory metals
    • C22B34/30—Obtaining chromium, molybdenum or tungsten
    • C22B34/36—Obtaining tungsten
    • 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
    • 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
    • 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 present invention relates to the extraction of rare high melting point metal tungsten in the field of hydrometallurgy, and more particularly to a method for extracting tungsten from scheelite. Background technique
  • Tungsten is known as "industrial teeth” and "war metal”. Due to its low global reserves and strong demand, tungsten has long been included in the strategic reserve list in many countries, and its strategic position is irreplaceable. About two-thirds of the world's tungsten resources are in the form of scheelite.
  • the international treatment of scheelite mainly uses soda pressing method, which can stably decompose low-grade scheelite.
  • the slag contains wo 3 and is generally stable below 1%, but the amount of reagent is too large, generally reaching the theoretical amount of 3 Times. Laboratory research even takes 5-6 times.
  • the operating temperature is as high as 225 ° C, and the equipment pressure is 20 atm.
  • the acid decomposition method mainly uses hydrochloric acid to treat scheelite concentrate, and thermodynamic studies show that the reaction trend is very high.
  • hydrochloric acid decomposes tungstic acid is yellow gelatinously wrapped on the surface of undecomposed white tungsten, which tends to cause incomplete decomposition, and the acid corrosion and volatilization of hydrochloric acid is serious, and the working environment is bad.
  • tungstic acid is yellow gelatinously wrapped on the surface of undecomposed white tungsten, which tends to cause incomplete decomposition, and the acid corrosion and volatilization of hydrochloric acid is serious, and the working environment is bad.
  • a large amount of residual hydrochloric acid mother liquor is neutralized by lime.
  • the CaCl 2 solution is discharged and is currently largely discarded.
  • tungsten can form a heteropolyacid with a ratio of phosphorus to phosphorus, arsenic, silicon, etc., which is a heteropolyacid of 1:6 ⁇ 1:12
  • a small amount of phosphorus in the decomposition process of hydrochloric acid can make a large amount of tungsten dispersion and loss, so the hydrochloric acid decomposition process mainly deals with high-grade scheelite concentrate (requires low impurities such as phosphorus and arsenic) ). However, some people use it instead. In the leaching process, a small amount of phosphoric acid is added to make tungsten enter the solution. The problem of decomposition is affected by the tungstic acid encapsulation during the decomposition of hydrochloric acid. According to research (Liu Yu, Liu Qi, Lu Tiejun, You Dazhao.
  • This method provides phosphorus as a complexing agent for tungsten by the addition of phosphoric acid, calcium phosphate or phosphorite when decomposing with sulfuric acid, but requires special addition. Amount of NaCl to improve decomposition.
  • the sulfuric acid concentration of the solution is 241.5g/L, and then in the high concentration of strong acid sulfuric acid solution.
  • H. Razavizadeh has conducted detailed experiments in accordance with the conditions of the patent (Production of tungsten via leaching of scheelite with sulfuric acid. Minerals & Metallurgical Processing. 2006, 23(2): 67-72). In the case of NaCl, a better leaching effect can be obtained.
  • the decomposition rate of the concentrate with a grade of 74.7% W0 3 can reach nearly 96%, but the equivalent slag contains W0 3 still up to about 5%; and the treated grade is 59% W0.
  • the decomposition rate of 3 concentrates is only 62%, and the equivalent slag contains W0 3 up to about 20%!
  • the high concentration of phosphoric acid (P 2 0 5 concentration of 15% ⁇ 35 %) to decompose scheelite can greatly increase the formation rate of soluble phosphotungstic heteropoly acid, thereby avoiding the encapsulation of tungstic acid precipitation. (1 ). Moreover, phosphoric acid is less corrosive and has no volatilization problems like hydrochloric acid.
  • cooling crystallization or concentrated crystallization
  • the crystallized mother liquor can be used to return to the new round of leaching after the leaching agent is added to the initial level.
  • the object of the present invention is to provide a method for extracting tungsten from scheelite without pollution, low cost, low energy consumption, simple operation and high yield.
  • a method for extracting tungsten from scheelite firstly adding a mixed acid of sulfuric acid and phosphoric acid to a decomposition reaction tank, wherein the concentration of the mixed acid of sulfuric acid and phosphoric acid is H 2 S0 4 at a concentration of 150 g/L to 500 g/L, P 2 0 5 mass content is controlled at 15% ⁇ 35%; after heating to 70 ⁇ 100 °C, scheelite is added, the liquid-solid ratio is controlled at 3:1 ⁇ 8:1 L/kg, and the reaction is filtered for 1 ⁇ 6h.
  • the phosphotungstic acid crystal is obtained by crystallization, and the phosphotungstic acid crystal water is dissolved and converted into an ammonium tungstate solution for preparing APT, and the obtained crystal mother liquid is supplemented with phosphoric acid and water. Return to the leaching after the initial level.
  • the scheelite contains W0 3 by mass of 10% to 75% and a particle size of 150 um.
  • the filtrate is cooled to 30 to 50 ° C by cooling crystallization, or concentrated.
  • the method of crystallization is to concentrate the filtrate volume to 1/3 to 4/5 of the original volume, and then filter to obtain a phosphotungstic acid crucible.
  • Phosphotungstic acid crystals obtained crystals were dissolved in water to give containing W0 3 350 ⁇ 500g / L phosphotungstic acid solution using W0 3 200 ⁇ 300g / L of ammonium tungstate aqueous ammonia or transition ion exchange or solvent extraction to obtain a solution containing transition Solution.
  • the phosphorus content of the scheelite impurity is not strictly required. There is no need to set up a special phosphorus removal process in the scheelite ore dressing process, which saves the cost of phosphorus removal reagent and the loss of tungsten;
  • the use of high concentration of phosphoric acid is beneficial to reduce the supersaturation of calcium sulfate during the decomposition of tungsten ore. This is because the complexation of calcium ions by phosphoric acid causes the solubility of calcium sulfate to gradually increase with the increase of the concentration of phosphoric acid in the solution. It is possible to reduce the spontaneous nucleation rate of calcium sulfate, thereby promoting the formation of coarse crystals, and avoiding the formation of dense calcium sulfate film to hinder the decomposition reaction, achieving efficient atmospheric pressure leaching of scheelite, saving resources and Energy consumption, and its decomposition rate can reach more than 99%;
  • the invention realizes the recycling of phosphoric acid and sulfuric acid, and the P 2 0 5 loss can be reduced to less than 5% in the process; the sulfuric acid consumption is only the theoretical consumption of the Ca content in the mineral, which greatly reduces the leaching cost and the wastewater. Emissions;
  • DRAWINGS Figure 1 is a graph showing the solubility of phosphotungstic acid as a function of sulfuric acid concentration and temperature
  • Figure 2 is an XRD pattern of the phosphotungstic acid crystal obtained by the method of the present invention.
  • FIG. 3 is a process flow diagram of the present invention
  • Figure 4 is an XRD pattern of the decomposition slag of Example 1;
  • Figure 5 is an SEM image of the decomposition slag of Example 1;
  • Figure 6 is an XRD pattern of the decomposition slag of Comparative Example 1;
  • Figure 7 is an SEM image of the decomposition slag of Comparative Example 1.
  • Example 1 The invention is further illustrated by the following examples, without further limiting the invention.
  • Example 1 The invention is further illustrated by the following examples, without further limiting the invention.
  • the use of high concentration of phosphoric acid in the invention is beneficial for reducing the supersaturation of calcium sulfate during the decomposition of tungsten ore, because the phosphoric acid can complex calcium ions, and the solubility of calcium sulfate gradually increases with the increase of the concentration of phosphoric acid in the solution, and thus It is possible to reduce the spontaneous nucleation rate of calcium sulfate, thereby promoting the formation of coarse crystals, and avoiding the formation of dense calcium sulfate film to hinder the decomposition reaction, and achieving efficient atmospheric pressure leaching of scheelite.
  • the experimental results are as follows:
  • White tungsten ore (including WO 3 70.6%) lkg prepare a phosphoric acid solution with a P 2 0 5 content of 20%, then adjust the sulfuric acid to control the concentration of H 2 S0 4 at 150g / L, liquid to solid ratio 5: l L / kg , reaction temperature 90 ° C, reaction time 6 h.
  • the tungsten leaching rate was 99.3 %.
  • the XRD pattern and SEM image of the obtained decomposition slag are shown in Figs.
  • White tungsten ore (including WO 3 70.6%) lkg prepare a phosphoric acid solution with a P 2 0 5 content of 20%, then adjust sulfuric acid to control the concentration of H 2 S0 4 at 150g / L, liquid to solid ratio of 6: l L / kg , reaction temperature 80 ° C, reaction time 6 h.
  • the tungsten leaching rate was 99.2%.
  • the filtrate was filtered, and the filtrate was added to the consumed sulfuric acid, and the filtrate was concentrated and crystallized to a volume of 1/3 of the original volume. Under this condition, the crystallinity of the phosphotungstic acid was 85.3%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the phosphotungstic acid crystals were dissolved in water to obtain a phosphotungstic acid solution containing W0 3 489.3 g/L, and the solution was added to aqueous ammonia to prepare an ammonium tungstate solution containing W0 3 250.6 g/L.
  • Scheelite (including WO 3 70.6%) lkg prepare a phosphoric acid solution with a P 2 0 5 content of 15%, then adjust the sulfuric acid to control the concentration of H 2 S0 4 at 300 g / L, liquid to solid ratio 4: l L / kg , reaction temperature 90 ° C, reaction time 4 h.
  • the tungsten leaching rate was 99.5 %.
  • the filtrate was filtered, and the filtrate was charged with sulfuric acid, and then cooled to 30 ° C. Under these conditions, the crystallinity of the phosphotungstic acid was 82.1%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the thionic acid crystals were dissolved in water to obtain a phosphotungstic acid solution containing W0 3 396.7 g/L, and the solution was added to aqueous ammonia to prepare an ammonium tungstate solution containing W0 3 262.3 g/L.
  • White tungsten ore (including WO 3 70.6%) lkg prepare phosphoric acid solution with P 2 0 5 content of 35%, then adjust sulfuric acid to control H 2 S0 4 concentration at 200g/L, liquid to solid ratio 3: l L / kg , reaction temperature 70 ° C, reaction time 5 h.
  • the tungsten leaching rate was 99.0%.
  • the filtrate was filtered, and the filtrate was added to the sulfuric acid consumed, and the filtrate was concentrated and crystallized to a volume of 4/5 of the original volume. Under these conditions, the crystallinity of the phosphotungstic acid was obtained. 81.7%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the phosphotungstic acid crystal was dissolved in water to obtain a phosphotungstic acid solution containing W0 3 358.4 g/L.
  • the solution was adsorbed by D301 resin, and the tungsten adsorption rate was 99.1%.
  • the ammonia was hydrolyzed to obtain WO 3 209.3 g/L tungsten.
  • the ammonium acid solution, the ion exchange solution is used to redissolve the phosphotungstic acid crystals.
  • White tungsten ore (including W0 3 65.7%) lkg prepare phosphoric acid solution with P 2 0 5 content of 35%, then adjust sulfuric acid to control H 2 S0 4 concentration at 500g/L, liquid-solid ratio 8: l L/kg , reaction temperature 90 ° C, reaction time lh.
  • the tungsten leaching rate was 99.0%.
  • the filtrate was filtered, and the filtrate was charged with sulfuric acid, and then cooled to 50 C. Under these conditions, the crystallinity of the phosphotungstic acid was 62.7%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the phosphotungstic acid crystal was dissolved in water to obtain a phosphotungstic acid solution containing wo 3 404.1 g/L, and the solution was added to aqueous ammonia to prepare an ammonium tungstate solution containing W0 3 228.7 g/L. ,
  • Scheelite (including W0 3 65.7%) lkg prepare a phosphoric acid solution with a P 2 0 5 content of 25%, then adjust sulfuric acid to control the concentration of H 2 S0 4 at 250 g / L, liquid to solid ratio 4: l L / kg , reaction temperature 100 ° C, reaction time 3 h.
  • the tungsten leaching rate was 99.3 %.
  • the filtrate was filtered, and the filtrate was charged with sulfuric acid and then cooled to 40 ° C. Under this condition, the crystallinity of the phosphotungstic acid was 67.1%.
  • the phosphotungstic acid crystal was dissolved in water to obtain a phosphotungstic acid solution containing W0 3 425.8 g/L, and the solution was added to aqueous ammonia to prepare an ammonium tungstate solution containing W0 3 231.4 g/L.
  • the crystallization mother liquid obtained in Example 6 was added to sulfuric acid, phosphoric acid and water so that the P 2 0 5 mass fraction in the solution was 25%, and the H 2 S0 4 concentration was 250 g/L, which was used for decomposing scheelite.
  • Used scheelite (including W0 3 65.7 Lkg, liquid-solid ratio 4: l L / kg, reaction temperature 100 ° C, reaction time 3 h.
  • the tungsten leaching rate was 99.2%.
  • the filtrate was filtered, and the filtrate was added to the sulfuric acid to be concentrated, and the filtrate was concentrated to a volume of 1/2 of the original volume. Under the conditions, the crystallinity of the phosphotungstic acid was 86.4%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the phosphotungstic acid crystal is dissolved in water to obtain a phosphotungstic acid solution containing W0 3 367.3 g/L, and the solution is added to ammonia water to obtain W0 3 253.8 g/L, P 2.4 g/L, S0 4 2 - 25.6 g. /L ammonium tungstate solution.
  • the ammonium magnesium salt method is used for purifying and removing impurities.
  • the MgCl 2 solution (MgCl 2 200g/L) is added in a molar ratio of Mg/P of 1.2, and the reaction is carried out for 30 minutes at room temperature, and the phosphorus removal rate is 99.9%, and the tungsten loss is only 0.05. %.
  • the molybdenum removal is then carried out using the method disclosed in patent 97108113.1.
  • the solution after removal of molybdenum is crystallized by evaporation to obtain APT.
  • the crystallization rate of APT is 94.5%, and the analysis results of the products are shown in Table 1.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • the phosphotungstic acid crystals were dissolved in water to obtain a phosphotungstic acid solution containing W0 3 376.8 g/L.
  • the solution was extracted with a primary amine-based alkaline extractant, and the extraction rate was 98.8%.
  • the extraction was carried out with ammonia water to obtain WO 3 205.2. g/L ammonium tungstate solution
  • the liquid, raffinate is used to redissolve the crystals of phosphotungstic acid.
  • White tungsten ore (including WO 3 30.4%) lkg prepare a phosphoric acid solution with a P 2 0 5 content of 25%, then adjust the sulfuric acid to control the concentration of H 2 S0 4 at 500 g / L, liquid to solid ratio 3: l L / kg , reaction temperature 90 ° C, reaction time 4 h.
  • the tungsten leaching rate was 98.9%.
  • the filtrate was filtered, and the filtrate was added to the sulfuric acid to be concentrated, and the filtrate was concentrated to a volume of 1/3 of the original volume. Under these conditions, the crystallinity of the phosphotungstic acid was 72.9%.
  • the mother liquor after filtration of the crystals is replenished with phosphoric acid and water to the initial level and returned to the next round of leaching.
  • Phosphotungstic acid was dissolved in water to give crystals containing W0 3 364.7g / L phosphotungstic acid solution, aqueous ammonia solution was added to the prepared containing W0 3 225.1g / L of ammonium tungstate solution.

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Description

一种从白钨矿中提取钨的方法 技术领域 本发明涉及湿法冶金领域中稀有高熔点金属钨的提取， 具体来说是一种 从白钨矿中提取钨的方法。 背景技术
钨被誉为 "工业牙齿"、 "战争金属"， 由于全球储量少、 需求强劲， 钨在 许多国家早就被列入战略储备清单， 其战略地位不可替代。 世界上约有 2/3的钨资源是以白钨矿形式存在。 国际上处理白钨矿主要采 用苏打压煮法， 该方法可以稳定地分解低品位白钨矿， 渣含 wo3—般稳定在 1%以下， 但是试剂用量太大， 一般要达到理论量的 3倍。有实验室研究甚至要 用到 5-6倍。 而且操作温度也高达 225°C左右， 设备压力达 20atm。 而在中国目前还没有形成专门的白钨矿处理工艺， 主要是延用黑钨矿 NaOH压煮的处理技术来处理白钨矿 (该工艺产量占 80%以上)。 该工艺在高碱 浓度， 高碱用量， 高温高压的强化条件下使得白钨矿得以有效分解，渣含 W03 可降到 1-3%。 不过这也造成了该工艺能耗大， 生产成本高， 后续工序废水排 放量大等缺点， 不符合我国节能减排、 低碳经济的发展要求。
酸分解法主要用盐酸来处理白钨精矿，热力学研究表明其反应趋势很高。 但盐酸分解时钨酸呈黄色胶状包裹在未分解的白钨表面， 容易导致分解不完 全， 而且盐酸的酸腐蚀和挥发问题严重， 工作环境恶劣， 最终大量的剩余盐 酸母液经石灰中和成 CaCl2溶液而排放， 目前已基本上弃置不用。
由于钨可与磷、 砷、 硅等杂质形成杂钨比为 1 :6〜1 :12的可溶性的杂多酸
(如 [PW1204Qf)， 在盐酸分解过程中少量的磷就可以造 i$大量钨的分散和损 失， 因此盐酸分解工艺主要处理高品位白钨精矿 (要求磷、 砷等杂质低)。 但 有人反其道而用之， 在浸出过程中特地加入少量磷酸使钨进入溶液， 从而解 决盐酸分解时的钨酸包裹而影响分解的问题。据研究 (刘玉， 刘琦， 卢铁军， 尤大钺. 几种不同白钨精矿盐酸络合浸取的动力学研究. 稀有金属与硬质合 金.1989,2: 10-16)报道， 在磷的用量少的时候仍会出现黄色钨酸， 因此需要 较大的过量系数， 而且磷的用量越大浸出速度越快。 但是盐酸的腐蚀和挥发 问题仍然令人头痛， 所以虽有研究报道但未见工业应用。
为了克服上述方法中盐酸的腐蚀和挥发问题， 可考虑采用硫酸来替代盐 酸， 以实现钨的络合浸出。 但是在大量硫酸存在时， 溶液对硫酸钙的过饱和 度极大， 往往导致石膏迅速成核形成大量细密结晶而造成了包裹， 使得分解 效果不理想。 如美国专利 4168296报道了用硫酸来浸出白钨矿的研究， 该方 法在采用硫酸分解时， 通过加入磷酸、 磷酸钙或磷灰岩以提供磷作为钨的络 合剂， 但是还需要特别加入一定量的 NaCl 以改善分解。' 据该专利描述， 为 分解 100g品位 29.6%W03的钨矿， 用 80g硫酸和 300ml水配成浸出液， 则 折合溶液硫酸浓度为 241.5g/L， 再在此高浓度的强酸性硫酸溶液中加入 4gNaCl0这时强酸性的硫酸溶液中 HC1的活度将会很高， 其实就相当于使用 了高浓度的盐酸， 其挥发问题和 CI—腐蚀问题又会出现。 发明者添加了 4g磷 灰石矿以供反应产生磷酸来结合钨， 磷灰石即便按纯 Ca5(P04)3F计算， 全部 分解后折合溶液中 P205含量仅为 0.3%， 而前已提及磷酸用量少时容易生成 钨酸沉淀。 这样石膏和钨酸两者的共同包裹作用严重阻碍了分解的进行。 因 此， 该分解方法仍需要采用三段浸出才能达到较高分解率。
实际上， H. Razavizadeh 曾按照该专利的条件进行了详细的实验 (Production of tungsten via leaching of scheelite with sulfuric acid. Minerals & Metallurgical Processing. 2006, 23(2): 67-72), 发现在添加了 NaCl的情况下才 可获得较好的浸出效果， 品位为 74.7% W03的精矿的分解率可达到近 96%， 但折合渣含 W03仍高达约 5%; 而处理品位为 59% W03的精矿时的分解率 仅有近 62%， 折合渣含 W03更高达约 20% ! 并且我们按该专利的方法， 使 用不同品位、 不同产地的白钨矿实验也发现， 一段浸出的分解率大约在 70-80%之间， 过滤也十分困难。 实验证明也确实需要经三段逆流浸出， 还要 将来自上段的分解渣再磨以破坏包裹， 分解率才可以达到 97-98%。
至于添加 NaCl的作用机理， 据文献 (Calcium Sulfate Dihydrate Nucleation in the Presence of Calcium and Sodium Chloride Salts. Ind. Eng. Chem. Res. 2001， 40， 2335-2339),氯离子的存在可较大幅度地增加硫酸钙结晶的诱导期， 从而阻滞自发成核。 这当然在一定程度上有利于形成粗大晶体， 避免产物包 裹矿物而影响分解。 但实际加入氯化钠仍难以达到满意的分解效果， 而且引 入 CT还会带来设备腐蚀和 HC1挥发的问题， 也不能解决钨酸包裹的问题。
采用高浓度的磷酸（P205浓度为 15 %〜35 % )来分解白钨矿可大幅提高 可溶性的磷钨杂多酸的形成速率， 从而避免了生成钨酸沉淀的包裹作用， 反 应式 （1 )。 而且磷酸的腐蚀性较低， 也没有像盐酸那样的挥发问题。
12CaW04+25H3P04= 12Ca(H2P04)2. ¾O+H3[PW12O40]+l lH2O (1) 但是磷酸的成本较高， 大量的磷以 Ca(H2P04:)2. H20形式存在难以进一 步利用。 为此， 在分解过程中可配入一定量的 H2S04 与生成的 Ca(H2P04)2 · H20进一步反应， 而使钙以硫酸钙形式沉淀并重新生成磷酸， 以降低磷酸的消耗， 反应式如 (2)。
Ca(H2P04)2. H20+H2S04→CaS04'n¾0+H3P04 (2) 总的反应式为
12CaW04+H3P04+ 12H2S04+ 12nH20= 12CaS04-nH20+H3 [P W12O40]+ 12H20 (3) 另外， 高浓度的磷酸有利于降低钨矿分解时硫酸钙的过饱和度， 这是由 于磷酸可络合钙离子， 导致硫酸钙的溶解度随着溶液中磷酸浓度的增加而逐 步上升， 大约在 P205为 20%左右达到最大 （80°C时， 为无磷酸存在时的 5-7 倍。 即便 P205升高到 40%左右， 溶解度也仍可达无磷酸存在时的 3-5倍）。 进 而也就有可能降低硫酸钙的自发成核速率， 从而起到促进形成粗大晶体的作 用， 而避免致密硫酸钙膜的生成来阻碍分解反应。 为此， 采用高浓度磷酸来分解白钨矿， 并配入一定量的硫酸来再生磷酸 的思路， 既可解决钨酸包裹问题， 使得钨以磷钨杂多酸的形式进入到溶液中 去， 又可避免产生致密硫酸钙膜， 从而达到高效分解白钨矿的目的。 另外， 经研究发现， 磷钨酸的溶解度受磷酸、 硫酸浓度和反应温度的影响比较显著， 这一特性为磷钨酸与浸出剂的分离提供了思路 （比如， 在磷酸浓度为含 P205 20%的体系下，我们实验测得的磷钨酸的溶解度随硫酸浓度和温度的变化图为 图 1， 结晶得到的晶体为图 2)。 即在高的酸浓度下可采用冷却结晶 （或浓缩 结晶） 使磷钨酸从浸出液中结晶出来， 从而实现钨与浸出剂的分离， 这样将 极大简化工艺的流程。 结晶后的母液补入浸出剂到初始水平后可用于返回新 一轮浸矿。
发明内容
本发明的目的是提供一种无污染， 成本低， 能耗低， 操作简单， 收率高 的从白钨矿中提取钨的方法。
本发明的目的是通过以下方式实现的：
一种从白钨矿中提取钨的方法： 首先将硫酸和磷酸的混合酸加入到分解 反应槽中， 所用的硫酸和磷酸混合酸浓度为 H2S04浓度在 150g/L〜500g/L， P205质量含量控制在 15%〜35%；升温至 70〜100°C后加入白钨矿，液固比控制 在 3:1~8:1 L/kg, 反应 l〜6h后过滤， 得到的滤液补入反应所消耗的硫酸后釆 用结晶的方法得到磷钨酸晶体， 磷钨酸结晶水溶解后经转型得到钨酸铵溶液 用于制备 APT, 得到的结晶母液补入磷酸和水到初始水平后返回浸矿。
所述的白钨矿含 W03质量百分比为 10%〜75 %， 粒径 150um。
上述方法中采用冷却结晶的方法将滤液冷却至 30〜50°C过滤，或者采用浓 缩结晶的方法将滤液体积浓縮至原来体积的 1/3〜4/5， 然后过滤得到磷钨酸结 曰
曰曰
结晶得到的磷钨酸晶体经水溶解后得到含 W03 350〜500g/L 的磷钨酸溶 液，采用氨水转型或离子交换或溶剂萃取来转型得到含 W03 200〜300g/L的钨 酸铵溶液。
本发明具有的优点是：
1.所处理的白钨矿杂质磷含量没有严格要求，白钨矿选矿过程中不必再设 专门除磷工序， 节约了除磷试剂成本及钨的损失；
2. 采用高浓度的磷酸有利于降低钨矿分解时硫酸钙的过饱和度， 这是由 于磷酸可络合钙离子， 导致硫酸钙的溶解度随着溶液中磷酸浓度的增加而逐 步上升， 进而也就有可能降低硫酸钙的自发成核速率， 从而起到促进形成粗 大晶体的作用， 而避免致密硫酸钙膜的生成来阻碍分解反应， 实现了白钨矿 的高效常压浸出， 节省了资源和能源消耗， 而且其分解率可达 99%以上；
3.在高浓度的磷酸和硫酸浸出体系下，浸出得到的磷钨酸的溶解度随硫酸 浓度和温度变化比较显著， 因而可采用冷却结晶或浓缩续晶的简单方式而实 现钨的提取。
4.克服了传统的酸分解工艺中的 Cf腐蚀严重问题和 HC1挥发严重问题；
5.本发明实现了磷酸和硫酸的循环利用， 过程中 P205 损失可降低到 5%以 下； 硫酸消耗量仅为矿物中 Ca含量的理论消耗量， 极大降低了浸出成本和废 水的排放量;
6.整个流程短， 操作方便， 易于实现工业化。
附图说明 图 1 为磷钨酸溶解度随硫酸浓度和温度的变化图；
图 2 为本发明方法得到的磷钨酸结晶的 XRD图；
图 3 为本发明的工艺流程图；
图 4 为实施例 1分解渣的 XRD图；
图 5 为实施例 1分解渣的 SEM图；
图 6 为对比例 1分解渣的 XRD图；
图 7 为对比例 1分解渣的 SEM图。
具体实施方式
下面结合实施例进一步说明本发明， 而不是对本发明的进一步限定。 实施例 1
本发明采用高浓度的磷酸有利于降低钨矿分解时硫酸钙的过饱和度，这是 由于磷酸可络合钙离子， 导致硫酸钙的溶解度随着溶液中磷酸浓度的增加而 逐步上升， 进而也就有可能降低硫酸钙的自发成核速率， 从而起到促进形成 粗大晶体的作用， 而避免致密硫酸钙膜的生成来阻碍分解反应， 实现了白钨 矿的高效常压浸出， 实验结果如下：
白钨矿 （含 WO370.6% ) lkg, 配制 P205含量为 20%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 150g/L， 液固比 5:l L / kg， 反应温度 90°C， 反应时 间 6h。 钨浸出率为 99.3 %。 得到的分解渣的XRD图和SEM图如图4, 5所示。
对比例 1
而采用较低磷酸浓度的条件下， 实验结果如下：
白钨矿（含 WO370.6% ) lkg, 配制 P205含量为 5%的磷酸溶液， 然后调入 硫酸， 控制 H2S04浓度在 150g/L， 液固比 5:l L / kg， 反应温度 90°C， 反应时间 6h。 钨浸出率为 87.6%。 得到的分解渣的XRD图和SEM图如图6，7所示。
实施例 2
白钨矿 （含 WO370.6% ) lkg, 配制 P205含量为 20%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 150g/L， 液固比 6:l L / kg， 反应温度 80°C， 反应时 间 6h。钨浸出率为 99.2%。 反应结束后立即过滤， 滤液补入所消耗的硫酸后经 浓缩结晶将滤液体积浓縮至原来体积的 1/3， 在此条件下， 磷钨酸的结晶率为 85.3%。 过滤晶体后的母液补入磷酸和水到初始水平返回下一轮浸矿。 磷钨酸 晶体经水溶解后得到含 W03489.3g/L的磷钨酸溶液，该溶液加入至氨水中制得 含 W03250.6g/L的钨酸铵溶液。
实施例 3
白钨矿 （含 WO370.6% ) lkg, 配制 P205含量为 15%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 300g/L， 液固比 4:l L / kg， 反应温度 90°C， 反应时 间 4h。钨浸出率为 99.5 %。 反应结束后立即过滤， 滤液补入所消耗的硫酸后冷 却至 30°C， 在此条件下， 磷钨酸的结晶率为 82.1%。 过滤晶体后的母液补入磷 酸和水到初始水平返回下一轮浸矿。 磯钨酸晶体经水溶解后得到含 W03 396.7g/L的磷钨酸溶液， 该溶液加入至氨水中制得含 W03262.3g/L的钨酸铵溶 液。
实施例 4
白钨矿 （含 WO370.6% ) lkg, 配制 P205含量为 35%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 200g/L， 液固比 3: l L / kg， 反应温度 70°C， 反应时 间 5h。钨浸出率为 99.0 %。 反应结束后立即过滤， 滤液补入所消耗的硫酸后经 浓縮结晶将滤液体积浓缩至原来体积的 4/5， 在此条件下， 磷钨酸的结晶率为 81.7%。 过滤晶体后的母液补入磷酸和水到初始水平返回下一轮浸矿。 磷钨酸 晶体经水溶解后得到含 W03 358.4g/L的磷钨酸溶液， 该溶液采用 D301树脂吸 附钨， 钨吸附率为 99.1%， 经氨水解吸得到含 WO3 209.3g/L的钨酸铵溶液， 离 子交换后液用于重新溶解磷钨酸结晶。
实施例 5
白钨矿 （含 W0365.7% ) lkg, 配制 P205含量为 35%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 500g/L， 液固比 8: l L/ kg， 反应温度 90°C， 反应时 间 lh。钨浸出率为 99.0%。 反应结束后立即过滤， 滤液补入所消耗的硫酸后冷 却至 50 C， 在此条件下， 磷钨酸的结晶率为 62.7%。 过滤晶体后的母液补入磷 酸和水到初始水平返回下一轮浸矿。 磷钨酸晶体经水溶解后得到含 wo3 404.1g/L的磷钨酸溶液， 该溶液加入至氨水中制得含 W03228.7g/L的钨酸铵溶 液。、
实施例 6
白钨矿 （含 W0365.7% ) lkg, 配制 P205含量为 25%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 250g/L， 液固比 4:l L / kg， 反应温度 100°C， 反应 时间 3h。 钨浸出率为 99.3 %。 反应结束后立即过滤， 滤液补入所消耗的硫酸后 冷却至 40°C， 在此条件下， 磷钨酸的结晶率为 67.1%。 磷钨酸晶体经水溶解后 得到含 W03425.8g/L的磷钨酸溶液， 该溶液加入至氨水中制得含 W03231.4g/L 的钨酸铵溶液。
实施例 7
将实施例 6中得到的结晶母液补入硫酸、 磷酸和水， 使得溶液中 P205质量 分数为 25%, H2S04浓度为 250g/L， 用于分解白钨矿。所用白钨矿（含 W0365.7 ) lkg, 液固比 4:l L / kg， 反应温度 100°C， 反应时间 3h。钨浸出率为 99.2%。 反应结束后立即过滤， 滤液补入所消耗的硫酸后浓缩结晶将滤液体积浓縮至 原来体积的 1/2， 在此条件下， 磷钨酸的结晶率为 86.4%。过滤晶体后的母液补 入磷酸和水到初始水平返回下一轮浸矿。 磷钨酸晶体经水溶解后得到含 W03 367.3g/L的磷钨酸溶液， 该溶液加入至氨水中制得含 W03253.8 g/L, P 2.4 g/L, S04 2—25.6 g/L的钨酸铵溶液。采用铵镁盐法进行净化除杂， MgCl2溶液（MgCl2 200g/L) 加入量为 Mg/P的摩尔比 1.2, 室温下反应 30min后过滤， 除磷率达到 99.9%， 钨损仅为 0.05%。然后采用专利 97108113.1公布的方法进行除钼。除钼 后的溶液经蒸发结晶得到 APT。 APT的结晶率为 94.5%， 其产品的分析结果见 表 1.
表 1结晶率为 94.5%的 APT的分析结果
元素 含量 （％) 元素 含量 （％) 元素 含量 （％)
P 0.00043 Mg <0.0007 Bi <0.0001
K <0細 Ni <0.0005 Sn <0.0001
Na 0.001 Ti <0.0005 Sb 0.0002
Mo 0.0034 V <0.0005 Cu <0.0001
Al <0.0005 Co <0.0005 Ca <0.0005
Si <0.0005 Cd Cr <0.0010
Mn <0.0005 Pb <0.0001 As 0.0007 实施例 8
白钨矿 （含 W0345.9% ) lkg, 配制 P205含量为 35%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 300g/L， 液固比 5:l L / kg， 反应温度 90°C， 反应时 间 6h。钨浸出率为 99.0%。 反应结束后立即过滤， 滤液补入所消耗的硫酸后浓 缩结晶将滤液体积浓缩至原来体积的 2/3， 在此条件下， 磷钨酸的结晶率为 84.7°/。。 过滤晶体后的母液补入磷酸和水到初始水平返回下一轮浸矿。 磷钨酸 晶体经水溶解后得到含 W03376.8g/L的磷钨酸溶液，该溶液采用伯胺基碱性萃 取剂萃取提取钨， 萃取率 98.8%， 用氨水反萃得到含 WO3 205.2g/L的钨酸铵溶 液， 萃余液用于重新溶解磷钨酸晶体。
实施例 9
白钨矿 （含 WO330.4% ) lkg, 配制 P205含量为 25%的磷酸溶液， 然后调 入硫酸， 控制 H2S04浓度在 500g/L， 液固比 3:l L / kg， 反应温度 90°C， 反应时 间 4h。钨浸出率为 98.9%。 反应结束后立即过滤， 滤液补入所消耗的硫酸后浓 缩结晶将滤液体积浓縮至原来体积的 1/3， 在此条件下， 磷钨酸的结晶率为 72.9%。 过滤晶体后的母液补入磷酸和水到初始水平返回下一轮浸矿。 磷钨酸 晶体经水溶解后得到含 W03364.7g/L的磷钨酸溶液，该溶液加入至氨水中制得 含 W03225.1g/L的钨酸铵溶液。

Claims

权 利 要 求
1、 一种从白钨矿中提取钨的方法， 其特征在于： 首先将硫酸和磷酸的混 合酸加入到分解反应槽中， 所用的硫酸和磷酸混合酸浓度为 H2S04浓度在 150g/L~500g/L， P205质量含量控制在 15%〜35%； 升温至 70〜100°C后加入白 钨矿， 液固比控制在 3:1〜8:1 L/kg, 反应 l〜6h后过滤， 得到的滤液补入反应 所消耗的硫酸后采用结晶的方法得到磷钨酸晶体， 磷钨酸结晶水溶解后经转 型得到钨酸铵溶液用于制备 APT, 得到的结晶母液补入磷酸和水到初始水平 后返回浸矿。
2、 根据权利要求 1所述的方法， 其特征在于:所述的白钨矿含 W03质量 百分比为 10%〜75%， 粒径 150腿。
3、 根据权利要求 1所述的方法， 其特征在于： 采用冷却结晶的方法将滤 液冷却至 30〜50°C过滤，或者采用浓缩结晶的方法将滤液体积浓缩至原来体积 的 1/3~4/5， 然后过滤得到磷钨酸结晶。
4、 根据权利要求 1或 3所述的方法， 其特征在于： 结晶得到的磷钨酸晶 体经水溶解后得到含 W03 350~500g/L的磷钨酸溶液，采用氨水转型或离子交 换或溶剂萃取来转型得到含 W03 200〜300g/L的钨酸铵溶液。
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