CN115232960B - Treatment method of mixed rare earth concentrate and application of quartz - Google Patents
Treatment method of mixed rare earth concentrate and application of quartz Download PDFInfo
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- CN115232960B CN115232960B CN202210869691.3A CN202210869691A CN115232960B CN 115232960 B CN115232960 B CN 115232960B CN 202210869691 A CN202210869691 A CN 202210869691A CN 115232960 B CN115232960 B CN 115232960B
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- 229910052761 rare earth metal Inorganic materials 0.000 title claims abstract description 177
- 150000002910 rare earth metals Chemical class 0.000 title claims abstract description 141
- 239000012141 concentrate Substances 0.000 title claims abstract description 114
- 238000000034 method Methods 0.000 title claims abstract description 65
- 239000010453 quartz Substances 0.000 title claims abstract description 58
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N silicon dioxide Inorganic materials O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 title claims abstract description 58
- 238000010438 heat treatment Methods 0.000 claims abstract description 27
- 239000002253 acid Substances 0.000 claims abstract description 26
- KRHYYFGTRYWZRS-UHFFFAOYSA-M Fluoride anion Chemical compound [F-] KRHYYFGTRYWZRS-UHFFFAOYSA-M 0.000 claims abstract description 14
- 238000002156 mixing Methods 0.000 claims abstract description 13
- IKNAJTLCCWPIQD-UHFFFAOYSA-K cerium(3+);lanthanum(3+);neodymium(3+);oxygen(2-);phosphate Chemical compound [O-2].[La+3].[Ce+3].[Nd+3].[O-]P([O-])([O-])=O IKNAJTLCCWPIQD-UHFFFAOYSA-K 0.000 claims abstract description 6
- 229910052590 monazite Inorganic materials 0.000 claims abstract description 6
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 75
- 239000007787 solid Substances 0.000 claims description 51
- 238000002386 leaching Methods 0.000 claims description 50
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 46
- 239000007788 liquid Substances 0.000 claims description 45
- 239000002245 particle Substances 0.000 claims description 37
- 230000008569 process Effects 0.000 claims description 27
- 239000001488 sodium phosphate Substances 0.000 claims description 22
- 229910000162 sodium phosphate Inorganic materials 0.000 claims description 22
- RYFMWSXOAZQYPI-UHFFFAOYSA-K trisodium phosphate Chemical compound [Na+].[Na+].[Na+].[O-]P([O-])([O-])=O RYFMWSXOAZQYPI-UHFFFAOYSA-K 0.000 claims description 22
- 238000000926 separation method Methods 0.000 claims description 16
- 229910001122 Mischmetal Inorganic materials 0.000 claims description 15
- 239000002893 slag Substances 0.000 claims description 14
- 238000006243 chemical reaction Methods 0.000 claims description 13
- 238000009833 condensation Methods 0.000 claims description 3
- 230000005494 condensation Effects 0.000 claims description 3
- 239000004575 stone Substances 0.000 claims 1
- 229910052731 fluorine Inorganic materials 0.000 abstract description 38
- YCKRFDGAMUMZLT-UHFFFAOYSA-N Fluorine atom Chemical compound [F] YCKRFDGAMUMZLT-UHFFFAOYSA-N 0.000 abstract description 36
- 239000011737 fluorine Substances 0.000 abstract description 36
- 239000000243 solution Substances 0.000 description 38
- 238000011084 recovery Methods 0.000 description 24
- 229910052698 phosphorus Inorganic materials 0.000 description 23
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 description 21
- 239000011574 phosphorus Substances 0.000 description 21
- 235000011121 sodium hydroxide Nutrition 0.000 description 21
- 238000001914 filtration Methods 0.000 description 19
- 239000003513 alkali Substances 0.000 description 15
- 230000000052 comparative effect Effects 0.000 description 14
- 238000000227 grinding Methods 0.000 description 14
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 13
- 239000007789 gas Substances 0.000 description 11
- 238000005406 washing Methods 0.000 description 11
- PUZPDOWCWNUUKD-UHFFFAOYSA-M sodium fluoride Chemical compound [F-].[Na+] PUZPDOWCWNUUKD-UHFFFAOYSA-M 0.000 description 8
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Chemical compound O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 8
- 239000000463 material Substances 0.000 description 7
- -1 rare earth chloride Chemical class 0.000 description 7
- 239000000126 substance Substances 0.000 description 6
- 229910045601 alloy Inorganic materials 0.000 description 5
- 239000000956 alloy Substances 0.000 description 5
- TWRXJAOTZQYOKJ-UHFFFAOYSA-L Magnesium chloride Chemical compound [Mg+2].[Cl-].[Cl-] TWRXJAOTZQYOKJ-UHFFFAOYSA-L 0.000 description 4
- 230000009286 beneficial effect Effects 0.000 description 4
- 238000000605 extraction Methods 0.000 description 4
- 229910052500 inorganic mineral Inorganic materials 0.000 description 4
- 235000010755 mineral Nutrition 0.000 description 4
- 239000011707 mineral Substances 0.000 description 4
- 239000011775 sodium fluoride Substances 0.000 description 4
- 235000013024 sodium fluoride Nutrition 0.000 description 4
- 239000002699 waste material Substances 0.000 description 4
- 238000000354 decomposition reaction Methods 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 239000012065 filter cake Substances 0.000 description 3
- 238000004519 manufacturing process Methods 0.000 description 3
- 239000002351 wastewater Substances 0.000 description 3
- NLXLAEXVIDQMFP-UHFFFAOYSA-N Ammonia chloride Chemical compound [NH4+].[Cl-] NLXLAEXVIDQMFP-UHFFFAOYSA-N 0.000 description 2
- VTYYLEPIZMXCLO-UHFFFAOYSA-L Calcium carbonate Chemical compound [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 description 2
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 description 2
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 description 2
- 230000015572 biosynthetic process Effects 0.000 description 2
- 239000003575 carbonaceous material Substances 0.000 description 2
- 238000001816 cooling Methods 0.000 description 2
- 238000005265 energy consumption Methods 0.000 description 2
- 239000012535 impurity Substances 0.000 description 2
- 229910001629 magnesium chloride Inorganic materials 0.000 description 2
- ORUIBWPALBXDOA-UHFFFAOYSA-L magnesium fluoride Chemical compound [F-].[F-].[Mg+2] ORUIBWPALBXDOA-UHFFFAOYSA-L 0.000 description 2
- 229910001635 magnesium fluoride Inorganic materials 0.000 description 2
- 238000002844 melting Methods 0.000 description 2
- 230000008018 melting Effects 0.000 description 2
- 150000007522 mineralic acids Chemical class 0.000 description 2
- 239000000203 mixture Substances 0.000 description 2
- 239000002244 precipitate Substances 0.000 description 2
- 239000000047 product Substances 0.000 description 2
- 230000035484 reaction time Effects 0.000 description 2
- 238000004064 recycling Methods 0.000 description 2
- 229910052710 silicon Inorganic materials 0.000 description 2
- 239000010703 silicon Substances 0.000 description 2
- VEXZGXHMUGYJMC-UHFFFAOYSA-M Chloride anion Chemical compound [Cl-] VEXZGXHMUGYJMC-UHFFFAOYSA-M 0.000 description 1
- 238000003723 Smelting Methods 0.000 description 1
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 1
- 230000009471 action Effects 0.000 description 1
- 238000005904 alkaline hydrolysis reaction Methods 0.000 description 1
- 239000012670 alkaline solution Substances 0.000 description 1
- 229910000323 aluminium silicate Inorganic materials 0.000 description 1
- 235000019270 ammonium chloride Nutrition 0.000 description 1
- BFNBIHQBYMNNAN-UHFFFAOYSA-N ammonium sulfate Chemical compound N.N.OS(O)(=O)=O BFNBIHQBYMNNAN-UHFFFAOYSA-N 0.000 description 1
- 229910052921 ammonium sulfate Inorganic materials 0.000 description 1
- 235000011130 ammonium sulphate Nutrition 0.000 description 1
- 230000000903 blocking effect Effects 0.000 description 1
- KGBXLFKZBHKPEV-UHFFFAOYSA-N boric acid Chemical compound OB(O)O KGBXLFKZBHKPEV-UHFFFAOYSA-N 0.000 description 1
- 239000004327 boric acid Substances 0.000 description 1
- 229910000019 calcium carbonate Inorganic materials 0.000 description 1
- BVKZGUZCCUSVTD-UHFFFAOYSA-N carbonic acid Chemical compound OC(O)=O BVKZGUZCCUSVTD-UHFFFAOYSA-N 0.000 description 1
- 238000002425 crystallisation Methods 0.000 description 1
- 230000008025 crystallization Effects 0.000 description 1
- 238000006115 defluorination reaction Methods 0.000 description 1
- HNPSIPDUKPIQMN-UHFFFAOYSA-N dioxosilane;oxo(oxoalumanyloxy)alumane Chemical compound O=[Si]=O.O=[Al]O[Al]=O HNPSIPDUKPIQMN-UHFFFAOYSA-N 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 238000007654 immersion Methods 0.000 description 1
- 238000005259 measurement Methods 0.000 description 1
- 238000000691 measurement method Methods 0.000 description 1
- 239000011259 mixed solution Substances 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 230000007935 neutral effect Effects 0.000 description 1
- 238000006386 neutralization reaction Methods 0.000 description 1
- 238000010298 pulverizing process Methods 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 238000010025 steaming Methods 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 235000019640 taste Nutrition 0.000 description 1
- 230000009466 transformation Effects 0.000 description 1
- 239000002912 waste gas Substances 0.000 description 1
Classifications
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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
- C22B1/00—Preliminary treatment of ores or scrap
-
- 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
- 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
- C22B3/12—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic alkaline 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
- C22B59/00—Obtaining rare earth metals
-
- 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
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- Chemical & Material Sciences (AREA)
- Engineering & Computer Science (AREA)
- Mechanical Engineering (AREA)
- Geology (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Environmental & Geological Engineering (AREA)
- Materials Engineering (AREA)
- Life Sciences & Earth Sciences (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Geochemistry & Mineralogy (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The application discloses a method for treating mixed rare earth concentrate and application of quartz, wherein the method comprises the following steps: mixing mixed rare earth concentrate and quartz, heating to 200-400 ℃, introducing steam, continuously heating to 850-950 ℃, reacting at the temperature, and condensing and recovering fluoride gas to obtain fluosilicic acid and defluorinated rare earth concentrate; wherein the mass ratio of quartz to the mixed rare earth concentrate is 0.04-0.06:1, and the mixed rare earth concentrate mainly comprises monazite and bastnaesite. The treatment method can obtain higher fluorine removal rate at lower treatment temperature.
Description
Technical Field
The application relates to a method for treating mixed rare earth concentrate and application of quartz.
Background
The smelting process of the mixed rare earth concentrate mainly comprises a sulfuric acid roasting method and a caustic soda method. The sulfuric acid roasting is divided into two processes of low-temperature concentrated sulfuric acid roasting and high-temperature concentrated sulfuric acid roasting. Mixed gas such as HF and SO generated in the process of roasting and decomposing concentrate by concentrated sulfuric acid 3 、SO 2 CO and CO 2 The method is difficult to treat such as CO, water vapor, sulfuric acid vapor and the like, and the whole process involves sulfuric acid roasting, water leaching, alkaline solution neutralization, extraction transformation, hydrochloric acid stripping and the like, and finally the rare earth chloride solution is obtained to carry out multistage extraction grouping of rare earth. The sulfuric acid roasting process has the problems of long flow, various raw materials, large consumption, high cost, large production amount of three wastes, high treatment difficulty, difficult effective separation and recovery of a large amount of phosphorus elements in concentrate into rare earth mixed solution and the like, and causes the waste of phosphorus and fluorine resources. The technological process of caustic soda method includes rare earth concentrate-decalcification treatmentCaustic soda decomposition, filtering and washing, hydrochloric acid dissolution, rare earth chloride mixing, multistage extraction and the like. The process has the problems of high rare earth concentrate grade requirement, long process flow, large washing water consumption, difficult treatment of a large amount of sodium fluoride and sodium phosphate entering washing liquid, and the like.
Therefore, a process capable of fully recovering fluorine, phosphorus, rare earth and other resources in mixed rare earth concentrate and simultaneously reducing the generation of wastewater, waste residue and waste gas is needed.
CN106586992a discloses a process for comprehensively recovering fluorine and phosphorus by liquid alkali decomposition of mixed rare earth concentrate: mixing the high-grade mixed rare earth concentrate with sodium hydroxide solution with the concentration of more than 60wt percent according to the weight ratio of the mixed rare earth concentrate to the sodium hydroxide of 1:3.5-7.5, reacting for 0.2-1 h at 150-160 ℃, performing hot filtration to obtain concentrated alkali liquor and alkali cake, cooling the concentrated alkali liquor, and performing filtration to obtain a sodium phosphate product; mixing the alkali cake with water for washing, and filtering to obtain primary washing alkali liquor; the alkali cake is continuously washed to be neutral, and is dissolved by hydrochloric acid, and the pH value is controlled to obtain rare earth chloride solution; concentrating and filtering the primary washing alkali liquor to obtain a sodium fluoride product. Although the process can recover fluorine and phosphorus respectively, the fluorine removal rate and the phosphorus recovery rate in the process are still to be improved; the obtained sodium phosphate has high fluorine content and is difficult to remove in the subsequent steps; when partial sodium fluoride remains in the rare earth filter cake, the rare earth yield is reduced and the energy consumption is larger when the rare earth fluoride is produced by hydrochloric acid leaching process.
CN109837385a discloses a method for decomposing rare earth ore by heating, melting, converting and decomposing, adding carbon material into furnace, melting rare earth ore of endosperm dephosphorization and fluorine-fixing material by utilizing the functions of heating by carbonic acid material and generating arc heating, and using calcium carbonate as fluorine-fixing material. The fluorine-containing substance generated by the process is unfavorable for recycling, and the fluorine removal rate is still to be improved.
CN114480835a discloses a method for decomposing mixed rare earth concentrate, roasting and decomposing mixed rare earth concentrate, magnesium chloride and carbon powder under the action of microwaves to obtain roasting ore; leaching the roasted ore by adopting first inorganic acid to obtain acid leaching slag and a first rare earth solution; separating acid leaching slag to obtain magnesium fluoride and non-decomposed rare earth concentrate; alkali decomposing the non-decomposed rare earth concentrate to obtain alkali wastewater and alkali hydrolyzed ore; cooling, concentrating and crystallizing the alkali wastewater to obtain sodium phosphate and recovered alkali liquor; leaching the alkaline hydrolysis ore by adopting a second inorganic acid to obtain a second rare earth solution. The decomposition method can obtain magnesium fluoride, but the fluorine removal rate is still to be improved.
Disclosure of Invention
In view of the above, an object of the present application is to provide a treatment method of mixed rare earth concentrates, which can obtain a higher fluorine removal rate at a lower treatment temperature. It is another object of the application to provide the use of quartz for treating misch metal concentrates to obtain fluosilicic acid. The application adopts the following technical scheme to realize the aim.
In one aspect, the application provides a method for treating mixed rare earth concentrate, comprising the following steps:
mixing mixed rare earth concentrate and quartz, heating to 200-400 ℃, introducing steam, continuously heating to 850-950 ℃, reacting at the temperature, and condensing and recovering fluoride gas to obtain fluosilicic acid and defluorinated rare earth concentrate;
wherein the mass ratio of quartz to mixed rare earth concentrate is 0.04-0.06:1;
wherein the rare earth minerals in the mixed rare earth concentrate mainly comprise monazite and bastnaesite.
According to the treatment method of the application, the mass ratio of quartz to mixed rare earth concentrate is preferably 0.045-0.055:1.
According to the treatment method of the present application, it is preferable to continue heating to 850 to 950 ℃ and reacting at that temperature for 1 to 4 hours.
According to the treatment method of the present application, preferably, the mixed rare earth concentrate has a particle size of 200 mesh or less; the granularity of quartz is less than or equal to 200 meshes.
According to the treatment method of the present application, preferably, the temperature of condensation is 45 ℃ or less.
The treatment method according to the present application preferably further comprises the steps of:
leaching the defluorinated rare earth concentrate by using sodium hydroxide solution, and carrying out solid-liquid separation to obtain sodium phosphate solution and rare earth element-containing solid;
leaching the solid containing rare earth elements by hydrochloric acid, and carrying out solid-liquid separation to obtain rare earth leaching liquid and solid slag.
The treatment method according to the application preferably further comprises the following specific steps:
leaching the defluorinated rare earth concentrate with sodium hydroxide solution at 140-170 ℃ for 3-7 h, and carrying out solid-liquid separation to obtain sodium phosphate solution and rare earth element-containing solid;
leaching the solid containing rare earth elements with hydrochloric acid at 75-85 ℃ for 2-4 h, and carrying out solid-liquid separation to obtain rare earth leaching liquid and solid slag.
According to the treatment method of the present application, preferably, the concentration of the sodium hydroxide solution is 45 to 65wt%; the solid-liquid mass ratio of the defluorinated rare earth concentrate to the sodium hydroxide solution is 1 (0.8-1.5).
According to the treatment method of the present application, preferably, the concentration of hydrochloric acid is 4.5 to 7mol/L; the mass ratio of the solid containing rare earth elements to the solid-liquid of hydrochloric acid is 1: (3.5-6.5).
In another aspect, the application also provides a use of quartz for treating misch metal concentrate to obtain fluosilicic acid, comprising the steps of: mixing mixed rare earth concentrate and quartz, heating to 200-400 ℃, introducing steam, continuously heating to 850-950 ℃ and reacting at the temperature, and condensing and recovering fluoride gas to obtain fluosilicic acid and defluorinated rare earth concentrate;
wherein the mass ratio of quartz to mixed rare earth concentrate is 0.04-0.06:1;
wherein, the mixed rare earth concentrate mainly comprises monazite and bastnaesite.
The treatment method of the mixed rare earth concentrate can obtain higher fluorine removal rate at a lower treatment temperature. In addition, the treatment method can also recover the sodium phosphate solution, so that the phosphorus recovery rate is higher, and meanwhile, the higher rare earth recovery rate can also be ensured. The application adopts a shorter process flow to realize the recovery of fluorine, phosphorus and rare earth respectively. The application does not add any chloride (comprising magnesium chloride and ammonium chloride), ammonium sulfate, carbon powder or carbon material, boric acid and the like, thus improving the purity of the obtained fluosilicic acid and reducing the introduction of other impurities.
Detailed Description
The present application will be further described with reference to the following specific embodiments, but the scope of the present application is not limited thereto.
In one aspect, the application provides a method for treating a misch metal concentrate. In another aspect, the application provides the use of quartz for treating misch metal concentrates to obtain fluosilicic acid. In the above method or use, the steps are identical and thus are described in detail below. The treatment method or the application of the application comprises the steps of fluosilicic acid and defluorinated rare earth concentrate formation, alkali leaching and acid leaching. Optionally, the method further comprises the step of crushing and/or grinding the misch metal concentrate and/or quartz.
The mixed rare earth concentrate refers to mixed rare earth ore mainly composed of monazite and bastnaesite, such as baiyunebo mixed rare earth ore. Although the misch metal concentrate also contains bastnaesite, the treatment process of the misch metal concentrate and bastnaesite is significantly different due to the presence of other minerals such as monazite. If the same process is used to treat both, the treatment effect is very different.
< step of pulverization and/or grinding >
And respectively crushing and/or grinding the mixed rare earth concentrate and/or quartz to be treated to obtain mixed rare earth concentrate particles and quartz particles.
In certain embodiments, the misch metal concentrate is ground to a particle size of 200 mesh or less, preferably 200-230 mesh, to yield misch metal concentrate particles. In other embodiments, the quartz is ground to a particle size of 200 mesh or less, preferably 200 to 230 mesh, to yield quartz particles. Thus, the ore grinding energy consumption is reduced while the defluorination of the mixed rare earth concentrate is not influenced.
< fluosilicic acid and defluorinated rare earth concentrate formation step >
In the application, mixed rare earth concentrate and quartz are mixed, heated to a first temperature, steam is introduced, the mixture is continuously heated to a second temperature, and the mixture reacts at the second temperature, and fluoride gas is condensed and recovered at the same time, so that fluosilicic acid and defluorinated rare earth concentrate are obtained. Preferably, the mixed rare earth concentrate particles and the quartz particles are uniformly mixed, heated to a first temperature, introduced with steam, continuously heated to a second temperature, reacted at the temperature, and simultaneously condensed and recovered with fluoride gas to obtain fluosilicic acid and fluorine-removing rare earth concentrate.
The present application surprisingly found that by adding quartz and controlling the amount and reaction temperature thereof under the condition of introducing steam, the fluorine removal rate can be significantly improved and fluorosilicic acid with higher purity can be obtained.
In the application, mixed rare earth concentrate particles and quartz particles can be placed into a gas-solid reaction furnace for heating or reaction. The gas-solid reaction furnace can be a fluidized bed, a fluidized bed furnace and a rotary kiln, and is preferably a fluidized bed of the gas-solid reaction furnace.
The first temperature may be 200 to 400 ℃, preferably 250 to 350 ℃, more preferably 280 to 320 ℃. Then introducing steam, when the temperature is lower than 700 ℃, the steam flow is 0.5-1.5 m 3 Preferably 0.8 to 1.2m, in terms of steam flow rate per ton of ore per minute 3 /(t.min); when the temperature is higher than 700 ℃, the flow rate of the water vapor is 2-5 m 3 Preferably 3 to 4m 3 /(t.min). t represents ton.
The second temperature may be 850 to 950 ℃, preferably 850 to 930 ℃, more preferably 850 to 920 ℃, still more preferably 880 to 920 ℃, still more preferably 900 to 920 ℃. The reaction time may be 1 to 4 hours, preferably 1.5 to 3.5 hours, more preferably 2 to 3 hours. This is advantageous in improving the fluorine removal rate. The prior art can obtain higher fluorine removal rate at the temperature of more than 1000 ℃ generally, but the rare earth mineral dead burning phenomenon is easily caused by the overhigh temperature, so that the activity is obviously reduced. By adding quartz, the application can obtain higher fluorine removal rate at lower treatment temperature and ensure higher rare earth recovery rate.
The mass ratio of quartz to mixed rare earth concentrate is 0.04-0.06:1, preferably 0.045-0.055:1, more preferably 0.045-0.05:1, and even more preferably 0.045-0.048:1. Thus being beneficial to improving the fluorine removal rate and improving the content of the fluosilicic acid.
The condensation temperature is 45 ℃ or lower, preferably 40 ℃ or lower, and more preferably 30 ℃ or lower. Thus being beneficial to improving the recovery rate of fluorine and obtaining fluosilicic acid with higher yield.
After the reaction is finished, the heating system is firstly closed, then the steam system is closed, the temperature is reduced, and the high-quality fluorine-removing rare earth concentrate is recovered. The content of fluorine in the defluorinated rare earth concentrate is less than or equal to 0.45wt%.
The calculation formula of the fluorine removal rate is as follows: (amount of fluorine in misch metal concentrate-amount of fluorine in defluorinated rare earth concentrate)/amount of fluorine in misch metal concentrate x 100%. The fluorine removal rate of the present application is 95% or more, preferably 96% or more.
< alkaline Leaching step >
Leaching the defluorinated rare earth concentrate by using sodium hydroxide solution, and carrying out solid-liquid separation to obtain sodium phosphate solution and rare earth element-containing solid.
According to one embodiment of the application, the defluorinated rare earth concentrate is leached with sodium hydroxide solution at 140-170 ℃ for 3-7 h, and solid-liquid separation is carried out to obtain sodium phosphate solution and rare earth element-containing solid.
The concentration of the sodium hydroxide solution may be 45 to 65wt%, preferably 50 to 65wt%, more preferably 55 to 65wt%. The dosage of the sodium hydroxide solution is as follows: the solid-liquid mass ratio of the defluorinated rare earth concentrate to the sodium hydroxide solution is 1 (0.8-1.5), preferably 1 (1.0-1.5), and more preferably 1 (1.0-1.3). The leaching temperature may be 140 to 170 ℃, preferably 145 to 170 ℃, more preferably 150 to 160 ℃. The leaching time may be 3 to 7 hours, preferably 3.5 to 6 hours, more preferably 4.5 to 5.5 hours. Thus being beneficial to separating sodium phosphate from the solid containing rare earth elements and respectively improving the recovery rate of the sodium phosphate and the solid containing rare earth elements. And the consumption of sodium hydroxide solution is small, so that the waste liquid production can be obviously reduced.
The solid-liquid separation can be filtration, and a small amount of water can be adopted to wash the filter cake during filtration. The solid containing rare earth elements is mainly water-insoluble rare earth hydroxide and oxide, and the separation of rare earth and phosphorus elements is realized through filtration.
The purity of the sodium phosphate solution is more than 50%, the sodium phosphate solid with the purity of more than 98% can be obtained by subsequent alkali (NaOH) removal and purification crystallization, and the phosphorus recovery rate can be more than 86%, preferably more than or equal to 87%. The phosphorus recovery rate was calculated as follows: the amount of phosphorus element in the sodium phosphate solution/the amount of phosphorus element in the mixed rare earth concentrate is multiplied by 100%.
< acid leaching step >
Leaching the solid containing rare earth elements by hydrochloric acid, and carrying out solid-liquid separation to obtain rare earth leaching liquid and solid slag.
According to one embodiment of the application, the solid containing rare earth elements is leached with hydrochloric acid at 75-85 ℃ for 2-4 hours, and the solid-liquid separation is carried out to obtain rare earth leaching liquid and solid slag.
The concentration of hydrochloric acid may be 4.5 to 7mol/L, preferably 5 to 7mol/L, more preferably 5.5 to 6.5mol/L. The dosage of the hydrochloric acid is as follows: the mass ratio of the solid containing rare earth elements to the solid-liquid of hydrochloric acid is 1: (3.5 to 6.5), preferably 1: (4.0 to 6.0), more preferably 1: (4.5-5.5).
The leaching temperature with hydrochloric acid may be 75 to 85 ℃, preferably 78 to 85 ℃, more preferably 80 to 85 ℃. The leaching time may be 2 to 4 hours, preferably 2.2 to 3.5 hours, more preferably 2.5 to 3 hours. Thus being beneficial to improving the recovery rate of rare earth.
The solid-liquid separation method is not particularly limited, and filtration is preferable.
The rare earth leaching solution is a solution containing rare earth chloride. The rare earth leaching solution can be used for extraction grouping of rare earth elements after impurity removal. The solid slag is a solid containing aluminosilicate.
The rare earth recovery rate of the application is higher and is more than or equal to 92%, preferably more than or equal to 94%. The calculation formula of the rare earth recovery rate is as follows: the REO content in the rare earth leaching solution/REO content in the mixed rare earth concentrate is multiplied by 100 percent.
The method firstly removes fluorine, is favorable for subsequent leaching, does not generate a large amount of rare earth fluoride precipitate to enter the solid, and does not generate production problems such as fluoride precipitate blocking a pipeline and the like; in addition, the recovery rate of phosphorus is improved.
< measurement method >
The measurement is carried out in the national laboratory detection center for the research and comprehensive utilization of the Baiyunebo rare earth resources.
Determination of fluorine content: adopts a fluorine steaming method.
Determination of phosphorus element content: ICP method was used.
Determination of REO content: the plasma method and the volumetric method are adopted.
In the following examples and comparative examples, the water vapor flow rates were as follows: at a reaction temperature below 700 ℃, the steam flow is 0.5m 3 /(t.min); above 700 deg.C, the water vapor flow is 3m 3 /(t·min)。
Example 1
Grinding mixed rare earth concentrate (Bayan-obo mixed rare earth ore) to a granularity smaller than 200 meshes to obtain mixed rare earth concentrate particles; the chemical components of the alloy are REO 60.10%, sigma Fe 3.20%, F7.80%, P3.50%, caO 4.35%, baO 3.71%, S1.26% and ThO 2 :1.23％、Nb 2 O 5 0.06 percent and the other percent of 14.79 percent (mass percent).
Grinding quartz to a particle size of less than 200 meshes to obtain quartz particles.
Uniformly mixing mixed rare earth concentrate particles and quartz particles, heating to a first temperature of 300 ℃ in a fluidized bed of a gas-solid reaction furnace, introducing steam, continuously heating to a second temperature, reacting for 3 hours at the second temperature, and condensing and recycling fluoride gas to obtain fluosilicic acid and fluorine-removed rare earth concentrate; wherein the mass ratio of quartz to mixed rare earth concentrate is recorded as M.
Leaching the defluorinated rare earth concentrate by adopting a 60wt% sodium hydroxide solution (the solid-liquid mass ratio is 1:1.2) for 5 hours at 150 ℃, filtering and washing to obtain a sodium phosphate solution and a rare earth element-containing solid.
Leaching the solid containing rare earth elements with 6mol/L hydrochloric acid (solid-liquid mass ratio is 1:5) for 3 hours at 85 ℃, and filtering to obtain rare earth leaching liquid and solid slag.
Comparative example 1
The difference from example 1 is only that the condensed fluoride gas is directly absorbed with 0.5mol/L sodium hydroxide solution without adding quartz.
Comparative example 2
The only difference from example 1 is that the mass ratio M of quartz to misch metal concentrate is different.
Comparative example 3
The only difference from example 1 is that the second temperature is different.
Comparative example 4
The only difference from example 1 is that the second temperature is different.
Some of the process parameters for example 1 and comparative examples 1-4 and the results are shown in Table 1 below.
TABLE 1
| Numbering device | Example 1 | Comparative example 1 | Comparative example 2 | Comparative example 3 | Comparative example 4 |
| Second temperature/. Degree.C | 900 | 900 | 900 | 750 | 1200 |
| Reaction time/h | 3 | 3 | 3 | 3 | 3 |
| M | 0.045:1 | 0 | 0.08:1 | 0.045:1 | 0.045:1 |
| Fluorine removal Rate% | 95 | 70 | 93 | 55 | 99 |
| Phosphorus recovery% | 87 | 50 | 70 | 45 | 70 |
| Rare earth recovery rate% | 93 | 85 | 89 | 82 | 75 |
Note that: the mass ratio of quartz to mixed rare earth concentrate is recorded as M.
As is clear from comparison between example 1 and comparative example 1, when quartz is not added, the fluorine removal rate is significantly reduced, sodium fluoride in the immersion liquid is seriously out of standard, the fluorine removal process is increased, the phosphorus loss is caused, the recovery rate is reduced, and in addition, the recovery rate of rare earth is influenced by a small amount of fluorine brought into a filter cake.
As is clear from a comparison between example 1 and comparative example 2, if the amount of quartz added is too large, the silicon content in the sodium phosphate solution is large after alkaline leaching, and the silicon removal process needs to be increased, resulting in a decrease in the phosphorus recovery rate.
As is clear from the comparison between example 1 and comparative example 3, when the second temperature is too low, the fluorine removal rate is significantly reduced, and the fluorine removal process is increased, resulting in low phosphorus recovery rate and affecting the rare earth recovery rate.
As is clear from comparison between example 1 and comparative example 4, the second temperature is too high, and the dead burning phenomenon of the rare earth mineral is obvious although the fluorine removal rate is high, the activity is obviously reduced, and the recovery rate of the subsequent phosphorus and rare earth is greatly reduced.
Example 2
Grinding the mixed rare earth concentrate to a granularity smaller than 200 meshes to obtain mixed rare earth concentrate particles; the chemical components of the alloy are 65.10 percent of REO, 3.20 percent of Sigma Fe, 7.80 percent of F, 4.10 percent of P, 4.35 percent of CaO, 2.15 percent of BaO, 1.36 percent of S and ThO 2 :1.03％、Nb 2 O 5 0.06 percent and the other percent of 10.85 percent (mass percent).
Grinding quartz to a particle size of less than 200 meshes to obtain quartz particles.
Uniformly mixing mixed rare earth concentrate particles and quartz particles, heating the mixed materials to a first temperature of 300 ℃ in a fluidized bed of a gas-solid reaction furnace, introducing steam, continuously heating to a second temperature of 950 ℃ and reacting at the second temperature for 2.5 hours, and condensing and recovering fluoride gas to obtain fluosilicic acid and fluorine-removed rare earth concentrate; wherein the mass ratio M of quartz to mixed rare earth concentrate is 0.045:1.
Leaching the cooled defluorinated rare earth concentrate for 5 hours at 160 ℃ by adopting 55wt% sodium hydroxide solution (the solid-liquid mass ratio is 1:1), filtering and washing to obtain sodium phosphate solution and rare earth element-containing solid.
Leaching the solid containing rare earth elements with 7mol/L hydrochloric acid (solid-liquid mass ratio is 1:4) at 80 ℃ for 2.5h, and filtering to obtain rare earth leaching liquid and solid slag.
Example 3
Grinding the mixed rare earth concentrate to a granularity smaller than 200 meshes to obtain mixed rare earth concentrate particles; the chemical components of the alloy are 65.10 percent of REO, 3.20 percent of Sigma Fe, 7.80 percent of F, 4.10 percent of P, 4.35 percent of CaO, 2.15 percent of BaO, 1.36 percent of S and ThO 2 :1.03％、Nb 2 O 5 0.06 percent and the other percent of 10.85 percent (mass percent).
Grinding quartz to a particle size of less than 200 meshes to obtain quartz particles.
Uniformly mixing mixed rare earth concentrate particles and quartz particles, heating the mixed materials to a first temperature of 200 ℃ in a fluidized bed of a gas-solid reaction furnace, introducing steam, continuously heating to a second temperature of 950 ℃ and reacting at the second temperature for 2 hours, and condensing and recovering fluoride gas to obtain fluosilicic acid and fluorine-removed rare earth concentrate; wherein the mass ratio M of quartz to mixed rare earth concentrate is 0.045:1.
Leaching the cooled defluorinated rare earth concentrate for 3 hours at 170 ℃ by adopting a 65wt% sodium hydroxide solution (the solid-liquid mass ratio is 1:0.8), filtering and washing to obtain a sodium phosphate solution and a rare earth element-containing solid.
Leaching the solid containing rare earth elements with 7mol/L hydrochloric acid (solid-liquid mass ratio is 1:3.5) for 4 hours at 75 ℃, and filtering to obtain rare earth leaching liquid and solid slag.
Example 4
Grinding the mixed rare earth concentrate to a granularity smaller than 200 meshes to obtain mixed rare earth concentrate particles; the chemical components of the alloy are REO 52.05%, sigma Fe 3.90%, F6.80%, P3.10%, caO 6.37%, baO 5.80%, S1.57% and ThO 2 :1.15％、Nb 2 O 5 0.04 percent and the other percent of 19.22 percent (mass percent).
Grinding quartz to a particle size of less than 200 meshes to obtain quartz particles.
Uniformly mixing mixed rare earth concentrate particles and quartz particles, heating the mixed materials to a first temperature of 300 ℃ in a fluidized bed of a gas-solid reaction furnace, introducing steam, continuously heating to a second temperature of 950 ℃ and reacting at the second temperature for 2.5 hours, and condensing and recovering fluoride gas to obtain fluosilicic acid and fluorine-removed rare earth concentrate; wherein the mass ratio M of quartz to mixed rare earth concentrate is 0.05:1.
Leaching the cooled defluorinated rare earth concentrate for 5 hours at 170 ℃ by adopting a 60wt% sodium hydroxide solution (the solid-liquid mass ratio is 1:0.8), filtering and washing to obtain a sodium phosphate solution and a rare earth element-containing solid.
Leaching the solid containing rare earth elements with 6mol/L hydrochloric acid (solid-liquid mass ratio is 1:5) for 3 hours at 85 ℃, and filtering to obtain rare earth leaching liquid and solid slag.
Example 5
Grinding the mixed rare earth concentrate to a granularity smaller than 200 meshes to obtain mixed rare earth concentrate particles; the chemical components of the alloy are REO 52.05%, sigma Fe 3.90%, F6.80%, P3.10%, caO 6.37%, baO 5.80%, S1.57% and ThO 2 :1.15％、Nb 2 O 5 0.04 percent and the other percent of 19.22 percent (mass percent).
Grinding quartz to a particle size of less than 200 meshes to obtain quartz particles.
Uniformly mixing mixed rare earth concentrate particles and quartz particles, heating the mixed materials to a first temperature of 400 ℃ in a fluidized bed of a gas-solid reaction furnace, introducing steam, continuously heating to a second temperature of 900 ℃ and reacting at the second temperature for 2.5 hours, and condensing and recovering fluoride gas to obtain fluosilicic acid and fluorine-removed rare earth concentrate; wherein the mass ratio M of quartz to mixed rare earth concentrate is 0.05:1.
Leaching the cooled defluorinated rare earth concentrate for 7h at 140 ℃ by adopting 45wt% sodium hydroxide solution (the solid-liquid mass ratio is 1:1.5), filtering and washing to obtain sodium phosphate solution and rare earth element-containing solid.
Leaching the solid containing rare earth elements with 4.5mol/L hydrochloric acid (solid-liquid mass ratio is 1:6.5) for 2 hours at 85 ℃, and filtering to obtain rare earth leaching liquid and solid slag.
TABLE 2
| Numbering device | Example 2 | Example 3 | Example 4 | Example 5 |
| Fluorine removal Rate% | 97 | 95 | 98 | 95 |
| Phosphorus recovery% | 89 | 86 | 86 | 87 |
| Rare earth recovery rate% | 95 | 92 | 94 | 92 |
From examples 1 to 5, the treatment method of the application is applicable to mixed rare earth concentrates with different tastes.
The present application is not limited to the above-described embodiments, and any modifications, improvements, substitutions, and the like, which may occur to those skilled in the art, fall within the scope of the present application without departing from the spirit of the application.
Claims (10)
1. The method for treating the mixed rare earth concentrate is characterized by comprising the following steps of:
mixing mixed rare earth concentrate and quartz, heating to 200-320 ℃, introducing steam, continuously heating to 880-950 ℃, reacting at the temperature, and condensing and recovering fluoride gas to obtain fluosilicic acid and defluorinated rare earth concentrate;
wherein the mass ratio of quartz to mixed rare earth concentrate is 0.04-0.06:1;
wherein, the mixed rare earth concentrate mainly comprises monazite and bastnaesite.
2. The process according to claim 1, characterized in that the mass ratio of quartz to misch metal concentrate is 0.045-0.055:1.
3. The process according to claim 2, wherein the heating is continued to 880-950 ℃ and the reaction is carried out at this temperature for 1-4 hours.
4. The process of claim 1, wherein the misch metal concentrate has a particle size of 200 mesh or less; the granularity of quartz is less than or equal to 200 meshes.
5. The process of claim 1 wherein the temperature of condensation is less than or equal to 45 ℃.
6. The method according to any one of claims 1 to 5, further comprising the steps of:
leaching the defluorinated rare earth concentrate by using sodium hydroxide solution, and carrying out solid-liquid separation to obtain sodium phosphate solution and rare earth element-containing solid;
leaching the solid containing rare earth elements by hydrochloric acid, and carrying out solid-liquid separation to obtain rare earth leaching liquid and solid slag.
7. The process according to any one of claims 1 to 5, further comprising the specific steps of:
leaching the defluorinated rare earth concentrate with sodium hydroxide solution at 140-170 ℃ for 3-7 h, and carrying out solid-liquid separation to obtain sodium phosphate solution and rare earth element-containing solid;
leaching the solid containing rare earth elements with hydrochloric acid at 75-85 ℃ for 2-4 h, and carrying out solid-liquid separation to obtain rare earth leaching liquid and solid slag.
8. The process according to claim 7, wherein the concentration of the sodium hydroxide solution is 45 to 65wt%; the solid-liquid mass ratio of the defluorinated rare earth concentrate to the sodium hydroxide solution is 1 (0.8-1.5).
9. The method according to claim 7, wherein the concentration of hydrochloric acid is 4.5 to 7mol/L; the mass ratio of the solid containing rare earth elements to the solid-liquid of hydrochloric acid is 1: (3.5-6.5).
10. Use of quartz for the treatment of misch metal concentrates to obtain fluosilicic acid, characterized in that it comprises the following steps: mixing mixed rare earth concentrate and quartz, heating to 200-320 ℃, introducing steam, continuously heating to 880-950 ℃ and reacting at the temperature, and condensing and recovering fluoride gas to obtain fluosilicic acid and defluorinated rare earth concentrate;
wherein the mass ratio of quartz to mixed rare earth concentrate is 0.04-0.06:1;
wherein the mixed rare earth concentrate mainly comprises solitary stone and bastnaesite 。
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| RU2128377C1 (en) * | 1997-12-11 | 1999-03-27 | Институт экспериментальной минералогии РАН | Method for sintering concentrate of rare-earth elements |
| CN105385844A (en) * | 2015-10-27 | 2016-03-09 | 中国科学院过程工程研究所 | Roasting defluorination device and process |
| CN107475542A (en) * | 2017-07-17 | 2017-12-15 | 中国恩菲工程技术有限公司 | The method for handling rare earth ore concentrate |
| CN113025835A (en) * | 2020-07-28 | 2021-06-25 | 江西理工大学 | Method for efficiently extracting rare earth from bastnaesite |
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| RU2128377C1 (en) * | 1997-12-11 | 1999-03-27 | Институт экспериментальной минералогии РАН | Method for sintering concentrate of rare-earth elements |
| CN105385844A (en) * | 2015-10-27 | 2016-03-09 | 中国科学院过程工程研究所 | Roasting defluorination device and process |
| CN107475542A (en) * | 2017-07-17 | 2017-12-15 | 中国恩菲工程技术有限公司 | The method for handling rare earth ore concentrate |
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