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WO2015081775A1 - 一种综合利用高铬型钒钛磁铁精矿的方法 - Google Patents

一种综合利用高铬型钒钛磁铁精矿的方法 Download PDF

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Publication number
WO2015081775A1
WO2015081775A1 PCT/CN2014/089728 CN2014089728W WO2015081775A1 WO 2015081775 A1 WO2015081775 A1 WO 2015081775A1 CN 2014089728 W CN2014089728 W CN 2014089728W WO 2015081775 A1 WO2015081775 A1 WO 2015081775A1
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titanium
vanadium
chromium
slag
mass
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English (en)
French (fr)
Inventor
齐涛
王丽娜
赵龙胜
陈德胜
赵宏欣
于宏东
刘亚辉
曲景奎
薛天艳
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Institute of Process Engineering of CAS
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Institute of Process Engineering of CAS
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Priority claimed from CN201310654831.6A external-priority patent/CN103757199B/zh
Priority claimed from CN201310653270.8A external-priority patent/CN103757426B/zh
Application filed by Institute of Process Engineering of CAS filed Critical Institute of Process Engineering of CAS
Publication of WO2015081775A1 publication Critical patent/WO2015081775A1/zh
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    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03C—MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03C1/00—Magnetic separation
    • B03C1/005—Pretreatment specially adapted for magnetic separation
    • B03C1/015—Pretreatment specially adapted for magnetic separation by chemical treatment imparting magnetic properties to the material to be separated, e.g. roasting, reduction, oxidation
    • 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
    • C22B3/10—Hydrochloric acid, other halogenated 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
    • C22B34/00—Obtaining refractory metals
    • C22B34/20—Obtaining niobium, tantalum or vanadium
    • C22B34/22—Obtaining vanadium
    • 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
    • C22B34/32—Obtaining chromium
    • 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
    • C22B5/00—General methods of reducing to metals
    • C22B5/02—Dry methods smelting of sulfides or formation of mattes
    • C22B5/10—Dry methods smelting of sulfides or formation of mattes by solid carbonaceous reducing agents
    • 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
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the invention belongs to the field of hydrometallurgy, and in particular, the invention relates to a method for comprehensively utilizing a high chromium type vanadium titanium magnet concentrate.
  • Panxi, Sichuan is the largest vanadium-titanium magnetite base in China. Its TiO 2 reserves account for more than 35% of the world's titanium reserves, accounting for more than 90% of the proven reserves in China, mainly concentrated in Panzhihua, Baima, and Hongge. Hehe and four mining areas. The vanadium-titanium magnetite in the Hongge mining area is currently the largest vanadium-titanium magnetite deposit in China with a reserve of 3.545 billion tons.
  • the red, high-chromium vanadium-titanium magnetite concentrate Compared with the Panzhihua vanadium-titanium magnetite concentrate, the red, high-chromium vanadium-titanium magnetite concentrate has the same content of iron, titanium and vanadium, and the associated chromium content is higher.
  • the grade of chromium (Cr 2 O 3 0.5% to 2%) is 8 to 10 times that of Panzhihua vanadium-titanium magnetite concentrate.
  • blast furnace route there are two main methods for comprehensive utilization of high-chromium vanadium-titanium magnetite concentrates: blast furnace route and direct reduction route.
  • blast furnace route see Chinese patent CN101020970A
  • high-chromium vanadium-titanium magnetite concentrate is smelted in a blast furnace or an electric furnace to obtain molten iron containing chromium and vanadium.
  • the titanium-containing slag obtained by the method is difficult to recycle, which not only causes waste of resources, but also has the risk of causing environmental deterioration due to the presence of chromium.
  • the direct reduction route can be further divided into the “first vanadium after iron” method (see Chinese patent CN102061397A) and the “first iron after vanadium” method (see Chinese patent CN101082068A, CN101294242A).
  • the high chromium vanadium-titanium magnetite concentrate is mixed with the sodium salt and oxidized and calcined at a high temperature, and then the vanadium and chromium are leached with water, and the residue after the water immersion is mixed with the pulverized coal to make a ball. After that, direct reduction is carried out in a rotary hearth furnace or an electric furnace.
  • the method has long process flow, high energy consumption, and the added sodium salt decomposes at high temperature to release harmful gases and pollute the environment.
  • the high-chromium vanadium-titanium magnet concentrate is dried by the batching ball, and then directly reduced in a rotary kiln or a rotary hearth furnace, and then the obtained metallized pellet is charged into an electric furnace for melting. Separation.
  • the Chinese patent CN101082068A does not describe the process conditions of direct reduction and melting separation, and the Chinese patent CN101294242A adopts a three-step high-temperature process of rotary hearth furnace reduction-electric furnace melting separation-electric furnace oxygen smelting, which has complicated process and high energy consumption.
  • the existing patent literature on high chromium (Cr 2 O 3 0.5% to 2%) vanadium-titanium magnet concentrate is basically a method of high temperature reduction and vanadium chrome hot metal blowing, and the general direct reduction-melting process is The iron and titanium are completely separated, and the deep reduction method is generally adopted.
  • the existing titanium-rich slag preparation technology generally uses titanium concentrate as a raw material and directly smelts in an electric furnace to obtain a titanium-rich slag having a TiO 2 mass content of about 75%.
  • the technology is relatively mature, but it is not suitable for the smelting of direct reduction-electric furnace melting of titanium slag, which is mainly because the content of impurities in the titanium slag of the electric furnace is relatively high.
  • the method for preparing titanium-rich slag by using electric furnace melting titanium slag as raw material generally adopts hydrochloric acid leaching to improve the raw material after activation.
  • Common activation methods include sodium roasting, mechanical activation, and microwave strengthening. This is because the phase of the titanium slag in the electric furnace is very stable, and it is very difficult to extract directly using hydrochloric acid.
  • the object of the present invention is to provide a kind of industrial operation and low energy consumption for the shortcomings of low utilization rate of titanium, vanadium and chromium, high energy consumption and serious environmental pollution in the conventional blast furnace iron-converter steelmaking process.
  • the invention provides a method for preparing vanadium chromium titanium slag by using high chromium type vanadium titanium magnet concentrate, comprising the following steps:
  • step 2) partially reducing the mixture of step 1), the reduction temperature is 1000-1300 ° C, and the reduction time is 1-10 h to obtain a metallized material;
  • the metallized material obtained in the step 2) is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed to obtain iron fine powder and vanadium chromium titanium slag.
  • the high chromium type vanadium titanium magnet concentrate has a TFe mass content of more than 40%, a TiO 2 mass content of more than 9%, and a V 2 O 5 mass content. More than 0.4%, the mass content of Cr 2 O 3 is more than 0.5%; and the carbon-containing reducing agent in step 1) is preferably one or more of anthracite, bituminous coal, lignite, and coke.
  • the additive of the step 1) is any one or more of sodium carbonate, potassium carbonate, sodium tetraborate, sodium fluoride, and alkali metal sodium silicate.
  • the metallization ratio of the metallized material in the step 2) is 30% to 80%.
  • the magnetic separation strength of the magnetic separation in step 3) is 200 to 2000 Oersted.
  • the recovery rates of titanium, vanadium and chromium in the vanadium chromium titanium slag after magnetic separation are respectively greater than 93%; the mass content of TFe in the iron fine powder is greater than 90%; the TFe in the vanadium chromium titanium slag The mass content is less than 35%; the phase of the vanadium chromium titanium slag is more unstable, and the titanium slag obtained by melting at a high temperature Compared with, it is more conducive to efficient extraction of titanium resources.
  • the existing direct reduction-electric furnace melting process uses high-temperature melting process, high energy consumption, and the obtained titanium slag phase is very stable, and it is usually necessary to extract titanium resources after activation.
  • Common activation methods include sodium roasting, mechanical activation, and microwave strengthening. This not only greatly increases equipment investment, but also greatly increases the energy consumption of the process.
  • the present invention adopts a partial reduction-magnetic separation technique, which not only avoids the high-temperature melting process, but also controls the orientation of vanadium and chromium to be consistent with titanium, and the phase of the obtained vanadium-chromium-titanium slag is relatively unstable, which is easy to follow.
  • the treatment extracts titanium, vanadium and chromium.
  • the method for comprehensively utilizing a high chromium type vanadium titanium magnet concentrate of the present invention comprises the following steps:
  • step 2) partially reducing the mixture of step 1), the reduction temperature is 1000-1300 ° C, and the reduction time is 1-10 h to obtain a metallized material;
  • the metallized material obtained in the step 2) is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed to obtain iron fine powder and vanadium chromium titanium slag;
  • step 4) The intermediate slurry obtained in step 4) is subjected to solid-liquid separation to obtain leach residue and a leachate containing vanadium and chromium;
  • the leaching residue obtained in the step 5) is washed and desiliconized, and then dried at 100 to 200 ° C to obtain a titanium-rich slag.
  • the steps 1) to 3) are both specific steps of the above method for preparing vanadium chromium titanium slag, and further specifically, the high chromium type of the step 1)
  • the content of TFe of the vanadium-titanium magnetite concentrate is more than 40%, the mass content of TiO 2 is more than 9%, the mass content of V 2 O 5 is more than 0.4%, and the mass content of Cr 2 O 3 is more than 0.5%;
  • the carbonaceous reducing agent is preferably one or more of anthracite, bituminous coal, lignite, and coke.
  • the additive of the step 1) is any one or more of sodium carbonate, potassium carbonate, sodium tetraborate, sodium fluoride, and alkali metal sodium silicate.
  • the metallization material has a metallization ratio of 30% to 80%.
  • Step 3) The magnetic separation strength of the magnetic separation is 200 to 2000 Oersted.
  • the recovery rates of titanium, vanadium and chromium in the vanadium chromium titanium slag after magnetic separation are respectively greater than 93%; the mass content of TFe in the iron fine powder is greater than 90%; the TFe in the vanadium chromium titanium slag The mass content is less than 35%; the phase of the vanadium chromium titanium slag is more unstable, and is more favorable for the efficient extraction of titanium resources than the titanium slag obtained by high temperature melting.
  • the mass percentage concentration of the hydrochloric acid solution in the step 4) is 10% to 30%, and the leaching time is 1 to 10 hours.
  • the mass concentration of TFe in the vanadium-containing and chromium-containing acid leaching solution in step 5) is 10 to 50 g/L, and the mass concentration of V 2 O 5 is 1.0. ⁇ 4.5 g/L, the mass concentration of Cr 2 O 3 is 1.5 to 6.0 g/L, and the mass concentrations of TiO 2 and SiO 2 are both less than 1.5 g/L.
  • the titanium-rich slag of the step 6) has a mass content of TiO 2 of more than 75%.
  • the recovery of titanium in the method is greater than 98%, and the recovery rates of vanadium and chromium are both greater than 90%.
  • the invention also provides a method for preparing titanium-rich slag by using vanadium chromium titanium slag, comprising the following steps:
  • step 2) The intermediate slurry obtained in step 1) is subjected to solid-liquid separation to obtain leach residue and a leachate containing vanadium and chromium;
  • the method for preparing titanium-rich slag by using vanadium chromium titanium slag according to the invention wherein the vanadium chromium titanium slag has a TFe mass content of less than 35%, the TiO 2 mass content is 15% to 30%, and the V 2 O 5 mass content. It is 0.5% to 3.0%, and the mass content of Cr 2 O 3 is 0.5% to 3.0%.
  • the vanadium chromium titanium slag can be prepared by the method for preparing vanadium chromium titanium slag using the high chromium type vanadium titanium magnet concentrate provided by the above invention.
  • the mass percentage concentration of the hydrochloric acid solution in the step 1) is 10% to 30%; and the leaching time in the step 1) is preferably 1 to 10 hours.
  • the mass concentration of TFe in the vanadium-containing and chromium-containing acid leaching solution is 10 to 50 g/L, and the mass concentration of V 2 O 5 is 1.0. ⁇ 4.5 g/L, the mass concentration of Cr 2 O 3 is 1.5 to 6.0 g/L, and the mass concentrations of TiO 2 and SiO 2 are both less than 1.5 g/L.
  • the titanium-rich slag of the step 3) has a mass content of TiO 2 of more than 75%.
  • the recovery rate of titanium in the method for preparing titanium-rich slag by using vanadium-chromium-titanium slag is greater than 98%, and the recovery rates of vanadium and chromium are both greater than 90%.
  • the existing direct reduction-electric furnace melting process has high energy consumption due to the high-temperature melting process, and the obtained titanium
  • the slag phase is very stable and usually requires titanium resources for activation.
  • Common activation methods include sodium roasting, mechanical activation, and microwave strengthening. This not only greatly increases equipment investment, but also greatly increases the energy consumption of the process.
  • the present invention employs a partial reduction-magnetic separation-hydrochloric acid leaching technique, which not only avoids the high-temperature melting process, but also controls the orientation of vanadium and chromium to be consistent with titanium.
  • the phase of the obtained vanadium-chromium-titanium slag is relatively unstable, and is directly acid-immersed with hydrochloric acid. The extraction rate of vanadium and chromium is high, and the impurity removal rate is high.
  • the present invention first proposes a partial reduction technique, that is, a method of controlling the degree of reduction, which can realize the separation of iron from titanium, vanadium and chromium, and provides an effective way for comprehensive utilization of high chromium vanadium-titanium magnetite.
  • the recovery rates of titanium, vanadium and chromium are both higher than 93%, and the resource utilization rate is high.
  • the iron content of the iron fine powder of the present invention is more than 90%, and is a high-quality steelmaking raw material.
  • the vanadium chromium titanium slag of the present invention is obtained by magnetic separation. Compared with the titanium slag obtained by electric furnace melting, the mineral phase structure of vanadium chrome-titanium slag is more unstable. The acid leaching of vanadium chrome-titanium slag can not only greatly improve the grade of titanium slag, but also realize the separation of titanium and vanadium and chromium. Efficient separation greatly increases the recovery of titanium, vanadium and chromium.
  • the titanium-rich slag obtained by the invention has a TiO 2 content of more than 75%, and can be used as a sulfuric acid-based titanium white raw material, and a titanium-rich slag having a TiO 2 mass content of more than 88% can also be obtained as a chlorinated titanium white raw material.
  • the process for comprehensively utilizing high-chromium vanadium-titanium magnetite concentrate of the invention has mild reaction conditions and greatly improves resource utilization rate, wherein the recovery rate of titanium is greater than 98%, and the recovery rates of vanadium and chromium are greater than 90 %.
  • Embodiments 1 to 6 of the present invention are a process flow diagram of Embodiments 1 to 6 of the present invention.
  • Embodiments 7 to 12 of the present invention are a process flow diagram of Embodiments 7 to 12 of the present invention.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 55%, TiO 2 mass content is 12.1%, V 2 O 5 mass content is 0.53%, Cr 2 O 3 mass content is 1.10%)
  • the mixture of anthracite and sodium carbonate is mixed, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and sodium carbonate is 100:8:2.5; the obtained mixture is partially reduced at 1200 ° C, and the reduction time is 2 hours.
  • Metallization material with a metallization rate of 70% is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 600 Oe to obtain a TFe mass content of 93.5.
  • % iron fines and TFe, TiO 2 , V 2 O 5 and Cr 2 O 3 are 21.5%, 27.5%, 1.10% and 2.10% vanadium chromium titanium slag, respectively; the recovery of titanium in this process 96.5%, the recovery rate of vanadium is 94.2%, and the recovery rate of chromium is 93.8%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 55%, TiO 2 mass content is 12.1%, V 2 O 5 mass content is 0.53%, Cr 2 O 3 mass content is 1.10%)
  • the mixture of bituminous coal and sodium tetraborate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, bituminous coal and sodium tetraborate is 100:20:3; the obtained mixture is partially reduced at 1300 ° C, and the reduction time is 1 hour.
  • the metallization material having a metallization rate of 80% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed under a magnetic field strength of 1000 Oersted to obtain a mass content of TFe.
  • 95.5% of iron fines and TFA, TiO 2 , V 2 O 5 and Cr 2 O 3 are respectively 18.6%, 29.5%, 1.03% and 1.75% of vanadium chromium titanium slag; titanium recovery in this process The rate was 95.7%, the recovery of vanadium was 93.9%, and the recovery of chromium was 93.1%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 47.2%, TiO 2 mass content is 10.5%, V 2 O 5 mass content is 1.20%, Cr 2 O 3 mass content is 0.58%)
  • the lignite is mixed to prepare a mixed material, wherein the weight ratio of the vanadium-titanium magnet concentrate to the coke is 100 : 2; the obtained mixed material is partially reduced at 1000 ° C for a reduction time of 10 hours to obtain a metal having a metallization ratio of 30%.
  • the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed under a magnetic field strength of 2000 Oersted to obtain iron fine powder and TFe having a TFe mass content of 90.1%.
  • the content of TiO 2 , V 2 O 5 and Cr 2 O 3 was 34.8%, 15.5%, 1.52% and 0.80%, respectively.
  • the recovery of titanium in this process was 94.3%, and the recovery of vanadium was 94.3%. 92.1%, the recovery rate of chromium is 91.8%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe has a mass content of 40.5%, TiO 2 has a mass content of 9.2%, V 2 O 5 has a mass content of 0.86%, and Cr 2 O 3 has a mass content of 0.75%).
  • the mixture of anthracite and sodium silicate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and sodium silicate is 100:10:1; the obtained mixture is partially reduced at 1150 ° C, and the reduction time is 4 hours.
  • a metallization material having a metallization ratio of 62% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 200 Oersted to obtain a mass content of TFe.
  • 97.5% of iron fines and TFA, TiO 2 , V 2 O 5 and Cr 2 O 3 are respectively 18.8%, 20.5%, 1.21% and 1.07% of vanadium chromium titanium slag; titanium recovery in this process The rate was 93.2%, the recovery of vanadium was 91.5%, and the recovery of chromium was 90.9%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 47.2%, TiO 2 mass content is 10.5%, V 2 O 5 mass content is 1.20%, Cr 2 O 3 mass content is 0.58%)
  • the mixture is prepared by mixing coke and sodium silicate, wherein the weight ratio of vanadium-titanium magnetite concentrate, coke and sodium silicate is 100:6.5:4; the obtained mixture is partially reduced at 1200 ° C, and the reduction time is 2 hours.
  • the metallization material having a metallization rate of 76% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 800 Oersted to obtain a mass content of TFe.
  • 94.7% of iron fines and Tef, TiO 2 , V 2 O 5 and Cr 2 O 3 were respectively 17.8%, 25.4%, 2.04% and 1.17% of vanadium chromium titanium slag; titanium recovery in this process The rate was 92.8%, the recovery of vanadium was 90.6%, and the recovery of chromium was 90.3%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe has a mass content of 40.5%, TiO 2 has a mass content of 9.2%, V 2 O 5 has a mass content of 0.86%, and Cr 2 O 3 has a mass content of 0.75%).
  • the mixture of anthracite and sodium tetraborate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and additive is 100:4:10; the obtained mixture is partially reduced at 1000 ° C, and the reduction time is 10 hours.
  • Metallization material with a metallization rate of 45% is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 1600 Oersted to obtain a mass content of TFe of 92.1. % iron fines and TFe, TiO 2 , V 2 O 5 and Cr 2 O 3 have a mass content of 24.3%, 17.5%, 1.41% and 1.27% vanadium chromium titanium slag; the recovery of titanium in this process 95.8%, the recovery rate of vanadium is 93.9%, and the recovery rate of chromium is 93.5%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 55%, TiO 2 mass content is 12.1%, V 2 O 5 mass content is 0.53%, Cr 2 O 3 mass content is 1.10%)
  • the mixture of anthracite and sodium carbonate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and sodium carbonate is 100:8:2.5; the obtained mixture is partially reduced at 1200 ° C, and the reduction time is 2 hours.
  • Metallization material with a metallization rate of 70% is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 600 Oe to obtain a TFe mass content of 93.5.
  • the mass content of % iron fines and TFe, TiO 2 , V 2 O 5 and Cr 2 O 3 were 21.5%, 27.5%, 1.10% and 2.10% of vanadium chromium titanium slag, respectively.
  • the content of TFA, TiO 2 , V 2 O 5 and Cr 2 O 3 was 21.5%, 27.5%, 1.10% and 2.10% of vanadium chromium titanium slag mixed with 25% hydrochloric acid solution, and leached at 150 ° C.
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 4.5:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 36.5.
  • the mass concentration of g/L, V 2 O 5 is 2.5 g/L, the mass concentration of Cr 2 O 3 is 6.0 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 1.5 g/L; the obtained leaching residue is washed.
  • a titanium-rich slag having a TiO 2 mass content of 92.6% was obtained; in this process, the recovery of titanium was 98.6%, the recovery of vanadium was 94.5%, and the recovery of chromium was 95.4%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 55%, TiO 2 mass content is 12.1%, V 2 O 5 mass content is 0.53%, Cr 2 O 3 mass content is 1.10%)
  • the mixture of bituminous coal and sodium tetraborate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, bituminous coal and sodium tetraborate is 100:20:3; the obtained mixture is partially reduced at 1300 ° C, and the reduction time is 1 hour.
  • the metallization material having a metallization rate of 80% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed under a magnetic field strength of 1000 Oersted to obtain a mass content of TFe.
  • 95.5% of iron fine powder and Tef, TiO 2 , V 2 O 5 and Cr 2 O 3 have a mass content of 18.6%, 29.5%, 1.03% and 1.75% of vanadium chromium titanium slag;
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 10:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 16.7.
  • the mass concentration of g/L, V 2 O 5 is 1.2 g/L, the mass concentration of Cr 2 O 3 is 2.0 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 0.8 g/L; the obtained leaching residue is washed.
  • a titanium-rich slag having a TiO 2 mass content of 76% is obtained; in this process, the recovery of titanium is 98.1%, the recovery of vanadium is 90.5%, and the recovery of chromium is 90.2%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 47.2%, TiO 2 mass content is 10.5%, V 2 O 5 mass content is 1.20%, Cr 2 O 3 mass content is 0.58%)
  • the lignite is mixed to prepare a mixed material, wherein the weight ratio of the vanadium-titanium magnet concentrate to the coke is 100:2; the obtained mixed material is partially reduced at 1000 ° C for a reduction time of 10 hours to obtain a metal having a metallization ratio of 30%.
  • the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed under a magnetic field strength of 2000 Oersted to obtain iron fine powder and TFe having a TFe mass content of 90.1%.
  • the content of TiO 2 , V 2 O 5 and Cr 2 O 3 was 34.8%, 15.5%, 1.52% and 0.80% of vanadium chromium titanium slag, respectively.
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 3:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 46.2.
  • the mass concentration of g/L, V 2 O 5 is 4.1 g/L, the mass concentration of Cr 2 O 3 is 2.3 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 1.0 g/L; the obtained leaching residue is washed.
  • titanium-rich slag having a TiO 2 mass content of 80% is obtained; in this process, the recovery of titanium is 98.4%, the recovery of vanadium is 91.1%, and the recovery of chromium is 92.8%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe has a mass content of 40.5%, TiO 2 has a mass content of 9.2%, V 2 O 5 has a mass content of 0.86%, and Cr 2 O 3 has a mass content of 0.75%).
  • the mixture of anthracite and sodium silicate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and sodium silicate is 100:10:1; the obtained mixture is partially reduced at 1150 ° C, and the reduction time is 4 hours.
  • a metallization material having a metallization ratio of 62% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 200 Oersted to obtain a mass content of TFe.
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 7:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 14.8.
  • the mass concentration of g/L, V 2 O 5 is 1.9 g/L, the mass concentration of Cr 2 O 3 is 1.7 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 1.0 g/L; the obtained leaching residue is washed.
  • a titanium-rich slag having a mass content of TiO 2 of 89.7% was obtained; in this process, the recovery of titanium was 98.2%, the recovery of vanadium was 92.4%, and the recovery of chromium was 93.7%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe mass content is 47.2%, TiO 2 mass content is 10.5%, V 2 O 5 mass content is 1.20%, Cr 2 O 3 mass content is 0.58%)
  • the mixture is prepared by mixing coke and sodium silicate, wherein the weight ratio of vanadium-titanium magnetite concentrate, coke and sodium silicate is 100:6.5:4; the obtained mixture is partially reduced at 1200 ° C, and the reduction time is 2 hours.
  • the metallization material having a metallization rate of 76% is obtained; the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 800 Oersted to obtain a mass content of TFe.
  • the vanadium chromium titanium slag having a mass content of 94.7% of iron fines and TFe, TiO 2 , V 2 O 5 and Cr 2 O 3 was 17.8%, 25.4%, 2.04% and 1.17%, respectively.
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 5.5:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 21.5.
  • the mass concentration of g/L, V 2 O 5 is 3.9 g/L, the mass concentration of Cr 2 O 3 is 2.0 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 0.8 g/L; the obtained leaching residue is washed.
  • titanium-rich slag having a TiO 2 mass content of 90.3% is obtained; in this process, the recovery of titanium is 98.9%, the recovery of vanadium is 94.2%, and the recovery of chromium is 93.6%.
  • High chromium type vanadium-titanium magnetite concentrate (TFe has a mass content of 40.5%, TiO 2 has a mass content of 9.2%, V 2 O 5 has a mass content of 0.86%, and Cr 2 O 3 has a mass content of 0.75%).
  • the mixture of anthracite and sodium tetraborate is prepared, wherein the weight ratio of vanadium-titanium magnetite concentrate, anthracite and additive is 100:4:10; the obtained mixture is partially reduced at 1000 ° C, and the reduction time is 10 hours.
  • Metallization material with a metallization rate of 45% the obtained metallized material is crushed and ground to a particle size of 90% less than 0.074 mm, and magnetic separation is performed at a magnetic field strength of 1600 Oersted to obtain a mass content of TFe of 92.1.
  • the content of iron fines and TFe, TiO 2 , V 2 O 5 and Cr 2 O 3 was 24.3%, 17.5%, 1.41% and 1.27% of vanadium chromium titanium slag, respectively.
  • an intermediate slurry is obtained, wherein the liquid-solid mass ratio of the dilute hydrochloric acid to the leaching slag is 4:1; the intermediate slurry is subjected to solid-liquid separation to obtain a leaching slag and a leaching solution containing vanadium and chromium, and the mass concentration of TFe in the leaching solution is 29.2.
  • the mass concentration of g/L, V 2 O 5 is 2.8 g/L, the mass concentration of Cr 2 O 3 is 2.2 g/L, and the mass concentration of TiO 2 and SiO 2 are both less than 1.2 g/L; the obtained leaching residue is washed.
  • a titanium-rich slag having a TiO 2 mass content of 93.5% was obtained; in this process, the recovery of titanium was 98.6%, the recovery of vanadium was 92.6%, and the recovery of chromium was 91.8%.

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Abstract

一种综合利用高铬型钒钛磁铁精矿的方法，包括以下步骤：1）将高铬型钒钛磁铁精矿与含碳还原剂及添加剂混合制备混合物料；2）将混合物料进行部分还原，得到金属化物料；3）将金属化物料进行磁选分离，获得铁精粉和钒铬钛渣；4）将钒铬钛渣与盐酸溶液混合，浸取，得到中间浆料；5）将步骤4）得到的中间浆料进行固液分离，得到浸出渣和含钒、铬的浸出液；6）将步骤5）得到的浸出渣经洗涤、脱硅后，得到富钛渣。该方法首次提出采用部分还原与盐酸浸出相藕合的技术，不仅实现了铁与钛、钒和铬的分离，还实现了钛与钒、铬的高效分离。该方法反应条件温和，资源利用率高，且能得到高品位的铁精粉和富钛渣。

Description

一种综合利用高铬型钒钛磁铁精矿的方法 技术领域
本发明属于湿法冶金领域，具体地，本发明涉及一种综合利用高铬型钒钛磁铁精矿的方法。
背景技术
四川攀西地区是我国最大的钒钛磁铁矿基地，其TiO2储量占世界钛资源储量的35％以上，占国内已探明储量的90％以上，主要集中分布在攀枝花、白马、红格和太和四个矿区。红格矿区的钒钛磁铁矿是目前国内最大的钒钛磁铁矿矿床，储量高达35.45亿吨。与攀枝花钒钛磁铁精矿相比，红格高铬型钒钛磁铁精矿的铁、钛和钒元素的含量相当，且伴生的铬元素含量较高。铬的品位(Cr2O30.5％～2％)是攀枝花钒钛磁铁精矿的8～10倍。
目前，高铬型钒钛磁铁精矿的综合利用方法主要有两种：高炉路线和直接还原路线。在高炉路线中(见中国专利CN101020970A)，将高铬型钒钛磁铁精矿通过高炉或电炉冶炼获得含铬、钒的铁水。但该方法得到的含钛炉渣难以回收利用，不仅造成资源浪费，由于铬的存在还存在导致环境恶化的风险。直接还原路线又可以分为“先钒后铁”法(见中国专利CN102061397A)和“先铁后钒”法(见中国专利CN101082068A、CN101294242A)。在“先钒后铁”法中，高铬型钒钛磁铁精矿与钠盐混合并在高温下氧化焙烧，再用水浸取其中的钒、铬，水浸后的残渣与煤粉混合造球后，在转底炉或电炉中进行直接还原。该方法工艺流程长，能耗高，且加入的钠盐在高温下分解释放有害气体，污染环境。在“先铁后钒”法中，高铬型钒钛磁铁精矿经配料压球干燥后，进入回转窑或转底炉中进行直接还原，再将得到的金属化球团装入电炉进行熔化分离。但中国专利CN101082068A没有说明直接还原和熔化分离的工艺条件，而中国专利CN101294242A采用转底炉还原-电炉熔化分离-电炉吹氧冶炼三步高温过程，流程复杂，能耗高。
已有的关于高铬型(Cr2O30.5％～2％)钒钛磁铁精矿的专利文献基本上都是高温还原以及钒铬铁水吹炼的方法，而一般直接还原-熔分过程为使铁钛完全分离，普遍采用深度还原的方式。现有的富钛渣制备技术一般是以钛精矿为原料，直接在电炉中进行冶炼，得到TiO2质量含量约为75％的富钛渣。该技术较为成 熟，但不适用于冶炼直接还原-电炉熔分得到的钛渣，这主要是因为电炉熔分钛渣中的杂质含量较高。目前，以电炉熔分钛渣为原料制备富钛渣的方法一般都是将原料活化后进行盐酸酸浸提质。常见的活化手段有钠化焙烧、机械活化和微波强化等。这是因为电炉熔分钛渣物相很稳定，利用盐酸直接提取非常困难。未见有采用部分还原与盐酸浸出相耦合的技术处理高铬型钒钛磁铁精矿，在实现铁与钛、钒和铬分离，钛与钒、铬分离的同时制备富钛渣的专利或文献报道。
发明内容
本发明的目的是针对目前高炉炼铁-转炉炼钢传统工艺中钛、钒和铬资源利用率低、能耗高以及环境污染严重等缺点，提供一种具有工业操作性、能耗低的一种综合利用高铬型钒钛磁铁精矿的方法。
本发明提供了一种利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法，包括以下步骤：
1)将高铬型钒钛磁铁精矿与含碳还原剂及添加剂混合制备混合物料，其中，钒钛磁铁精矿∶含碳还原剂∶添加剂的质量比为100∶2～20∶0～10；
2)将步骤1)的混合物料进行部分还原，还原温度为1000～1300℃，还原时间为1～10h，得到金属化物料；
3)将步骤2)得到的金属化物料破碎、磨细至粒度90％小于0.074mm，进行磁选分离，获得铁精粉和钒铬钛渣。
根据本发明的制备钒铬钛渣的方法，步骤1)所述高铬型钒钛磁铁精矿的TFe的质量含量大于40％，TiO2的质量含量大于9％，V2O5的质量含量大于0.4％，Cr2O3的质量含量大于0.5％；步骤1)所述含碳还原剂优选为无烟煤、烟煤、褐煤、焦炭中的一种或几种。根据本发明的制备钒铬钛渣的方法，步骤1)所述的添加剂为碳酸钠、碳酸钾、四硼酸钠、氟化钠、硅酸钠碱金属盐中的任一种或几种。
根据本发明的制备钒铬钛渣的方法，步骤2)所述金属化物料的金属化率为30％～80％。
根据本发明的制备钒铬钛渣的方法，步骤3)所述磁选分离的磁场强度为200～2000奥斯特。磁选分离后的钒铬钛渣中钛、钒和铬的回收率分别均大于93％；所述的铁精粉中的TFe的质量含量大于90％；所述的钒铬钛渣中的TFe的质量含量小于35％；所述的钒铬钛渣的物相更不稳定，与高温熔分得到的钛渣 相比，更利于钛资源高效提取。
现有直接还原-电炉熔分流程由于采用高温熔分过程，能耗高，且得到的钛渣物相很稳定，通常需要进行活化后提取钛资源。常见的活化手段包括钠化焙烧、机械活化和微波强化等。这不仅大大增加了设备投资，还大大增加了工艺的能耗。针对这些问题，本发明采用部分还原-磁选分离技术，不仅避免了高温熔分过程，还能控制钒、铬的走向与钛一致，得到的钒铬钛渣的物相相对不稳定，易于后续处理提取钛、钒和铬。
本发明的综合利用高铬型钒钛磁铁精矿的方法，包括以下步骤：
1)将高铬型钒钛磁铁精矿与含碳还原剂及添加剂混合制备混合物料，其中，钒钛磁铁精矿∶含碳还原剂∶添加剂的质量比为100∶2～20∶0～10；
2)将步骤1)的混合物料进行部分还原，还原温度为1000～1300℃，还原时间为1～10h，得到金属化物料；
3)将步骤2)得到的金属化物料破碎、磨细至粒度90％小于0.074mm，进行磁选分离，获得铁精粉和钒铬钛渣；
4)将钒铬钛渣与盐酸溶液混合，在90～160℃下浸取，得到中间浆料；其中，所述的盐酸溶液与钒铬钛渣的液固质量比为3∶1～10∶1；
5)将步骤4)得到中间浆料进行固液分离，得到浸出渣和含钒、铬的浸出液；
6)将步骤5)得到的浸出渣经洗涤、脱硅后，在100～200℃下进行干燥，得到富钛渣。
根据本发明的综合利用高铬型钒钛磁铁精矿方法，所述步骤1)-3)既为上述制备钒铬钛渣的方法的具体步骤，进一步具体地，步骤1)所述高铬型钒钛磁铁精矿的TFe的质量含量大于40％，TiO2的质量含量大于9％，V2O5的质量含量大于0.4％，Cr2O3的质量含量大于0.5％；步骤1)所述含碳还原剂优选为无烟煤、烟煤、褐煤、焦炭中的一种或几种。根据本发明的制备钒铬钛渣的方法，步骤1)所述的添加剂为碳酸钠、碳酸钾、四硼酸钠、氟化钠、硅酸钠碱金属盐中的任一种或几种。步骤2)所述金属化物料的金属化率为30％～80％。步骤3)所述磁选分离的磁场强度为200～2000奥斯特。磁选分离后的钒铬钛渣中钛、钒和铬的回收率分别均大于93％；所述的铁精粉中的TFe的质量含量大于90％；所述的钒铬钛渣中的TFe的质量含量小于35％；所述的钒铬钛渣的物相更不稳定，与高温熔分得到的钛渣相比，更利于钛资源高效提取。
根据本发明的综合利用高铬型钒钛磁铁精矿方法，步骤4)所述的盐酸溶液的质量百分比浓度为10％～30％，所述的浸取时间为1～10小时。
根据本发明的综合利用高铬型钒钛磁铁精矿方法，步骤5)所述含钒、铬的酸浸液中TFe的质量浓度为10～50g/L，V2O5的质量浓度为1.0～4.5g/L，Cr2O3的质量浓度为1.5～6.0g/L，TiO2和SiO2的质量浓度均小于1.5g/L。
根据本发明的综合利用高铬型钒钛磁铁精矿方法，步骤6)所述的富钛渣中TiO2的质量含量大于75％。
根据本发明的综合利用高铬型钒钛磁铁精矿方法，所述方法中钛的回收率大于98％，钒和铬的回收率均大于90％。
本发明还提供了利用钒铬钛渣制备富钛渣的方法，包括以下步骤：
1)将钒铬钛渣与盐酸溶液混合，在90～160℃下浸取，得到中间浆料；其中，所述的盐酸溶液与钒铬钛渣的液固质量比为3∶1～10∶1；
2)将步骤1)得到中间浆料进行固液分离，得到浸出渣和含钒、铬的浸出液；
3)将步骤2)得到的浸出渣经洗涤、脱硅后，在100～200℃下进行干燥，得到富钛渣。
根据本发明的利用钒铬钛渣制备富钛渣的方法，所述钒铬钛渣的TFe的质量含量小于35％，TiO2的质量含量为15％～30％，V2O5的质量含量为0.5％～3.0％，Cr2O3的质量含量为0.5％～3.0％。
进一步地，所述钒铬钛渣可以使用上述本发明提供的利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法进行制备。
根据本发明的利用钒铬钛渣制备富钛渣的方法，步骤1)所述的盐酸溶液的质量百分比浓度为10％～30％；步骤1)所述的浸取时间优选为1～10小时。
根据本发明的利用钒铬钛渣制备富钛渣的方法，步骤2)所述含钒、铬的酸浸液中TFe的质量浓度为10～50g/L，V2O5的质量浓度为1.0～4.5g/L，Cr2O3的质量浓度为1.5～6.0g/L，TiO2和SiO2的质量浓度均小于1.5g/L。
根据本发明的利用钒铬钛渣制备富钛渣方法，步骤3)所述的富钛渣中TiO2的质量含量大于75％。
根据本发明的利用钒铬钛渣制备富钛渣的方法，所述的利用钒铬钛渣制备富钛渣的方法中钛的回收率大于98％，钒和铬的回收率均大于90％。
现有直接还原-电炉熔分流程由于采用高温熔分过程，能耗高，且得到的钛 渣物相很稳定，通常需要进行活化后提取钛资源。常见的活化手段包括钠化焙烧、机械活化和微波强化等。这不仅大大增加了设备投资，还大大增加了工艺的能耗。针对这些问题，本发明采用部分还原-磁选分离-盐酸浸出技术，不仅避免了高温熔分过程，还能控制钒、铬的走向与钛一致。得到的钒铬钛渣的物相相对不稳定，采用盐酸直接酸浸，钒、铬提取率高，杂质去除率高。
本发明的优点在于：
(1)本发明首次提出采用部分还原技术，即控制还原程度的方法，能实现铁与钛、钒和铬的分离，为综合利用高铬型钒钛磁铁矿提供了一条有效的途径。
(2)本发明的磁选分离中，钛、钒和铬的回收率均高于93％，资源利用率高。
(2)本发明的铁精粉的TFe的质量含量大于90％，是优质的炼钢原料。
(3)本发明的钒铬钛渣是通过磁选分离得到的。与电炉熔分得到的钛渣相比，钒铬钛渣的矿相结构更不稳定，采用盐酸酸浸钒铬钛渣，不仅能大幅提高钛渣的品位，还能实现钛与钒、铬的高效分离，大大提高钛、钒和铬的回收率。
(5)本发明得到的富钛渣的TiO2质量含量大于75％，可以作为硫酸法钛白原料，也可获得TiO2质量含量大于88％的富钛渣，作为氯化法钛白原料。
(6)本发明的综合利用高铬型钒钛磁铁精矿的工艺，反应条件温和，且大幅提高了资源利用率，其中，钛的回收率大于98％，钒、铬的回收率均大于90％。
附图说明
图1为本发明实施例1～6的工艺流程图。
图2为本发明实施例7～12的工艺流程图。
具体实施方式
实施例1
将高铬型钒钛磁铁精矿(TFe的质量含量为55％，TiO2的质量含量为12.1％，V2O5的质量含量为0.53％，Cr2O3的质量含量为1.10％)与无烟煤及碳酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和碳酸钠的重量比为100∶8∶2.5；得到的混合物料在1200℃下进行部分还原，还原时间为2小时，得到金属化率为70％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在600奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为93.5％的铁精 粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为21.5％、27.5％、1.10％和2.10％的钒铬钛渣；此工艺中钛的回收率为96.5％，钒的回收率为94.2％，铬的回收率为93.8％。
实施例2
将高铬型钒钛磁铁精矿(TFe的质量含量为55％，TiO2的质量含量为12.1％，V2O5的质量含量为0.53％，Cr2O3的质量含量为1.10％)与烟煤及四硼酸钠混合制备混合物料，其中，钒钛磁铁精矿、烟煤和四硼酸钠的重量比为100∶20∶3；得到的混合物料在1300℃下进行部分还原，还原时间为1小时，得到金属化率为80％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在1000奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为95.5％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.6％、29.5％、1.03％和1.75％的钒铬钛渣；此工艺中钛的回收率为95.7％，钒的回收率为93.9％，铬的回收率为93.1％。
实施例3
将高铬型钒钛磁铁精矿(TFe的质量含量为47.2％，TiO2的质量含量为10.5％，V2O5的质量含量为1.20％，Cr2O3的质量含量为0.58％)与褐煤混合制备混合物料，其中，钒钛磁铁精矿和焦炭的重量比为100∶2；得到的混合物料在1000℃下进行部分还原，还原时间为10小时，得到金属化率为30％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在2000奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为90.1％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为34.8％、15.5％、1.52％和0.80％的钒铬钛渣；此工艺中钛的回收率为94.3％，钒的回收率为92.1％，铬的回收率为91.8％。
实施例4
将高铬型钒钛磁铁精矿(TFe的质量含量为40.5％，TiO2的质量含量为9.2％，V2O5的质量含量为0.86％，Cr2O3的质量含量为0.75％)与无烟煤及硅酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和硅酸钠的重量比为100∶10∶1；得到的混合物料在1150℃下进行部分还原，还原时间为4小时，得到金属化率为62％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm， 在200奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为97.5％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.8％、20.5％、1.21％和1.07％的钒铬钛渣；此工艺中钛的回收率为93.2％，钒的回收率为91.5％，铬的回收率为90.9％。
实施例5
将高铬型钒钛磁铁精矿(TFe的质量含量为47.2％，TiO2的质量含量为10.5％，V2O5的质量含量为1.20％，Cr2O3的质量含量为0.58％)与焦炭及硅酸钠混合制备混合物料，其中，钒钛磁铁精矿、焦炭和硅酸钠的重量比为100∶6.5∶4；得到的混合物料在1200℃下进行部分还原，还原时间为2小时，得到金属化率为76％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在800奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为94.7％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为17.8％、25.4％、2.04％和1.17％的钒铬钛渣；此工艺中钛的回收率为92.8％，钒的回收率为90.6％，铬的回收率为90.3％。
实施例6
将高铬型钒钛磁铁精矿(TFe的质量含量为40.5％，TiO2的质量含量为9.2％，V2O5的质量含量为0.86％，Cr2O3的质量含量为0.75％)与无烟煤及四硼酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和添加剂的重量比为100∶4∶10；得到的混合物料在1000℃下进行部分还原，还原时间为10小时，得到金属化率为45％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在1600奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为92.1％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为24.3％、17.5％、1.41％和1.27％的钒铬钛渣；此工艺中钛的回收率为95.8％，钒的回收率为93.9％，铬的回收率为93.5％。
实施例7
将高铬型钒钛磁铁精矿(TFe的质量含量为55％，TiO2的质量含量为12.1％，V2O5的质量含量为0.53％，Cr2O3的质量含量为1.10％)与无烟煤及碳酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和碳酸钠的重量比为100∶8∶2.5； 得到的混合物料在1200℃下进行部分还原，还原时间为2小时，得到金属化率为70％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在600奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为93.5％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为21.5％、27.5％、1.10％和2.10％的钒铬钛渣。
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为21.5％、27.5％、1.10％和2.10％的钒铬钛渣与25％盐酸溶液混合，在150℃下浸取3小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为4.5∶1；中间浆料经固液分离得到浸出渣和含钒、铬的浸出液，浸出液中TFe的质量浓度为36.5g/L，V2O5的质量浓度为2.5g/L，Cr2O3的质量浓度为6.0g/L，TiO2和SiO2的质量浓度均小于1.5g/L；所得浸出渣经洗涤、脱硅、100℃干燥后得到TiO2的质量含量为92.6％的富钛渣；此工艺中钛的回收率为98.6％，钒的回收率为94.5％，铬的回收率为95.4％。
实施例8
将高铬型钒钛磁铁精矿(TFe的质量含量为55％，TiO2的质量含量为12.1％，V2O5的质量含量为0.53％，Cr2O3的质量含量为1.10％)与烟煤及四硼酸钠混合制备混合物料，其中，钒钛磁铁精矿、烟煤和四硼酸钠的重量比为100∶20∶3；得到的混合物料在1300℃下进行部分还原，还原时间为1小时，得到金属化率为80％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在1000奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为95.5％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.6％、29.5％、1.03％和1.75％的钒铬钛渣；
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.6％、29.5％、1.03％和1.75％的钒铬钛渣与10％盐酸溶液混合，在160℃下浸取1小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为10∶1；中间浆料经固液分离得到浸出渣和含钒、铬的浸出液，浸出液中TFe的质量浓度为16.7g/L，V2O5的质量浓度为1.2g/L，Cr2O3的质量浓度为2.0g/L，TiO2和SiO2的质量浓度均小于0.8g/L；所得浸出渣经洗涤、脱硅、150℃干燥后得到TiO2的质量含量为76％的富钛渣；此工艺中钛的回收率为98.1％，钒的回收率为90.5％，铬的回收率为90.2％。
实施例9
将高铬型钒钛磁铁精矿(TFe的质量含量为47.2％，TiO2的质量含量为10.5％，V2O5的质量含量为1.20％，Cr2O3的质量含量为0.58％)与褐煤混合制备混合物料，其中，钒钛磁铁精矿和焦炭的重量比为100∶2；得到的混合物料在1000℃下进行部分还原，还原时间为10小时，得到金属化率为30％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在2000奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为90.1％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为34.8％、15.5％、1.52％和0.80％的钒铬钛渣。
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为34.8％、15.5％、1.52％和0.80％的钒铬钛渣与30％盐酸溶液混合，在90℃下浸取10小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为3∶1；中间浆料经固液分离得到浸出渣和含钒、铬的浸出液，浸出液中TFe的质量浓度为46.2g/L，V2O5的质量浓度为4.1g/L，Cr2O3的质量浓度为2.3g/L，TiO2和SiO2的质量浓度均小于1.0g/L；所得浸出渣经洗涤、脱硅、200℃干燥后得到TiO2的质量含量为80％的富钛渣；此工艺中钛的回收率为98.4％，钒的回收率为91.1％，铬的回收率为92.8％。
实施例10
将高铬型钒钛磁铁精矿(TFe的质量含量为40.5％，TiO2的质量含量为9.2％，V2O5的质量含量为0.86％，Cr2O3的质量含量为0.75％)与无烟煤及硅酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和硅酸钠的重量比为100∶10∶1；得到的混合物料在1150℃下进行部分还原，还原时间为4小时，得到金属化率为62％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在200奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为97.5％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.8％、20.5％、1.21％和1.07％的钒铬钛渣。
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为18.8％、20.5％、1.21％和1.07％的钒铬钛渣与15％盐酸溶液混合，在120℃下浸取5小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为7∶1；中间浆料经固液分离得到浸出渣和含钒、铬的浸出液，浸出液中TFe的质量浓度为14.8g/L，V2O5的质量浓度为1.9g/L，Cr2O3的质量浓度为1.7g/L，TiO2和SiO2的质量浓度均小于1.0g/L；所得浸出渣经洗涤、脱硅、150℃干燥后得到TiO2的质量含量为89.7％的富钛渣；此工艺 中钛的回收率为98.2％，钒的回收率为92.4％，铬的回收率为93.7％。
实施例11
将高铬型钒钛磁铁精矿(TFe的质量含量为47.2％，TiO2的质量含量为10.5％，V2O5的质量含量为1.20％，Cr2O3的质量含量为0.58％)与焦炭及硅酸钠混合制备混合物料，其中，钒钛磁铁精矿、焦炭和硅酸钠的重量比为100∶6.5∶4；得到的混合物料在1200℃下进行部分还原，还原时间为2小时，得到金属化率为76％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在800奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为94.7％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为17.8％、25.4％、2.04％和1.17％的钒铬钛渣。
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为17.8％、25.4％、2.04％和1.17％的钒铬钛渣与20％盐酸溶液混合，在135℃下浸取4.5小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为5.5∶1；中间浆料经固液分离得到浸出渣和含钒、铬的浸出液，浸出液中TFe的质量浓度为21.5g/L，V2O5的质量浓度为3.9g/L，Cr2O3的质量浓度为2.0g/L，TiO2和SiO2的质量浓度均小于0.8g/L；所得浸出渣经洗涤、脱硅、100℃干燥后得到TiO2的质量含量为90.3％的富钛渣；此工艺中钛的回收率为98.9％，钒的回收率为94.2％，铬的回收率为93.6％。
实施例12
将高铬型钒钛磁铁精矿(TFe的质量含量为40.5％，TiO2的质量含量为9.2％，V2O5的质量含量为0.86％，Cr2O3的质量含量为0.75％)与无烟煤及四硼酸钠混合制备混合物料，其中，钒钛磁铁精矿、无烟煤和添加剂的重量比为100∶4∶10；得到的混合物料在1000℃下进行部分还原，还原时间为10小时，得到金属化率为45％的金属化物料；得到的金属化物料经破碎、磨细至粒度90％小于0.074mm，在1600奥斯特的磁场强度下进行磁选分离，获得TFe的质量含量为92.1％的铁精粉和TFe、TiO2、V2O5和Cr2O3的质量含量分别为24.3％、17.5％、1.41％和1.27％的钒铬钛渣。
将TFe、TiO2、V2O5和Cr2O3的质量含量分别为24.3％、17.5％、1.41％和1.27％的钒铬钛渣与25％盐酸溶液混合，在105℃下浸取9小时，得到中间浆料，其中，稀盐酸与浸出渣的液固质量比为4∶1；中间浆料经固液分离得到浸出渣和含钒、 铬的浸出液，浸出液中TFe的质量浓度为29.2g/L，V2O5的质量浓度为2.8g/L，Cr2O3的质量浓度为2.2g/L，TiO2和SiO2的质量浓度均小于1.2g/L；所得浸出渣经洗涤、脱硅、120℃干燥后得到TiO2的质量含量为93.5％的富钛渣；此工艺中钛的回收率为98.6％，钒的回收率为92.6％，铬的回收率为91.8％。
当然，本发明还可以有多种实施例，在不背离本发明精神及其实质的情况下，熟悉本领域的技术人员可根据本发明的公开做出各种相应的改变和变型，但这些相应的改变和变形都应属于本发明所附的权利要求的保护范围。

Claims (22)

  1. 一种利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法，包括以下步骤：
    1)将高铬型钒钛磁铁精矿与含碳还原剂及添加剂混合制备混合物料，其中，钒钛磁铁精矿：含碳还原剂：添加剂的质量比为100∶2～20∶0～10；
    2)将步骤1)的混合物料进行部分还原，还原温度为1000～1300℃，还原时间为1～10h，得到金属化物料；
    3)将步骤2)得到的金属化物料破碎、磨细至粒度90％小于0.074mm，进行磁选分离，获得铁精粉和钒铬钛渣。
  2. 根据权利要求1所述的制备钒铬钛渣的方法，其特征在于，步骤1)所述高铬型钒钛磁铁精矿的TFe的质量含量大于40％，TiO2的质量含量大于9％，V2O5的质量含量大于0.4％，Cr2O3的质量含量大于0.5％。
  3. 根据权利要求1所述的制备钒铬钛渣的方法，其特征在于，步骤1)所述含碳还原剂为无烟煤、烟煤、褐煤、焦炭中的一种或几种。
  4. 根据权利要求1所述的制备钒铬钛渣的方法，其特征在于，步骤1)所述的添加剂为碳酸钠、碳酸钾、四硼酸钠、氟化钠、硅酸钠碱金属盐中的一种或几种。
  5. 根据权利要求1所述的制备钒铬钛渣的方法，其特征在于，步骤2)所述金属化物料的金属化率为30％～80％。
  6. 根据权利要求1所述的制备钒铬钛渣的方法，其特征在于，步骤3)所述磁选分离的磁场强度为200～2000奥斯特。
  7. 根据权利要求1或6所述的制备钒铬钛渣的方法，其特征在于，步骤3)所述的磁选分离后的钒铬钛渣中钛、钒和铬的回收率分别均大于93％。
  8. 根据权利要求1或6所述的制备钒铬钛渣的方法，其特征在于，步骤3)所述的铁精粉中的TFe的质量含量大于90％。
  9. 根据权利要求1或6所述的制备钒铬钛渣的方法，其特征在于，步骤3)所述的钒铬钛渣中的TFe的质量含量小于35％。
  10. 一种综合利用高铬型钒钛磁铁精矿的方法，包括以下步骤：
    1)将高铬型钒钛磁铁精矿与含碳还原剂及添加剂混合制备混合物料，其中，钒钛磁铁精矿：含碳还原剂：添加剂的质量比为100∶2～20∶0～10；
    2)将步骤1)的混合物料进行部分还原，还原温度为1000～1300℃，还原时间为1～10h，得到金属化物料；
    3)将步骤2)得到的金属化物料破碎、磨细至粒度90％小于0.074mm，进行磁选分离，获得铁精粉和钒铬钛渣；
    4)将钒铬钛渣与盐酸溶液混合，在90～160℃下浸取，得到中间浆料；其中，所述的盐酸溶液与钒铬钛渣的液固质量比为3∶1～10∶1；
    5)将步骤4)得到中间浆料进行固液分离，得到浸出渣和含钒、铬的浸出液；
    6)将步骤5)得到的浸出渣经洗涤、脱硅后，在100～200℃下进行干燥，得到富钛渣。
  11. 根据权利要求10所述的方法，其特征在于，步骤4)所述的盐酸溶液的质量百分比浓度为10％～30％。
  12. 根据权利要求10所述的方法，其特征在于，步骤4)所述的浸取时间为1～10小时。
  13. 根据权利要求10所述的方法，其特征是：步骤5)所述含钒、铬的酸浸液中TFe的质量浓度为10～50g/L，V2O5的质量浓度为1.0～4.5g/L，Cr2O3的质量浓度为1.5～6.0g/L，TiO2和SiO2的质量浓度均小于1.5g/L。
  14. 根据权利要求10所述的方法，其特征在于，步骤6)所述的富钛渣中TiO2的质量含量大于75％。
  15. 根据权利要求10所述的方法，其特征在于，所述方法中钛的回收率大于98％，钒和铬的回收率均大于90％。
  16. 一种利用钒铬钛渣制备富钛渣的方法，包括以下步骤：
    1)将钒铬钛渣与盐酸溶液混合，在90～160℃下浸取，得到中间浆料；其中，所述的盐酸溶液与钒铬钛渣的液固质量比为3∶1～10∶1；
    2)将步骤1)得到中间浆料进行固液分离，得到浸出渣和含钒、铬的浸出液；
    3)将步骤2)得到的浸出渣经洗涤、脱硅后，在100～200℃下进行干燥，得到富钛渣。
  17. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征在于，所述钒铬钛渣的TFe的质量含量小于35％，TiO2的质量含量为15％～35％，V2O5的质量含量为0.5％～3.0％，Cr2O3的质量含量为0.5％～3.0％。
  18. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征是：步骤1)所述的盐酸溶液的质量百分比浓度为10％～30％。
  19. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征是：步骤1)所述的浸取时间为1～10小时。
  20. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征是：步骤2)所述含钒、铬的酸浸液中TFe的质量浓度为10～50g/L，V2O5的质量浓度为1.0～4.5g/L，Cr2O3的质量浓度为1.5～6.0g/L，TiO2和SiO2的质量浓度均小于1.5g/L。
  21. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征在于，步骤3)所述的富钛渣中TiO2的质量含量大于75％。
  22. 根据权利要求16所述利用钒铬钛渣制备富钛渣的方法，其特征在于，所述方法中钛的回收率大于98％，钒和铬的回收率均大于90％。
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