CN114480881A - Method for extracting valuable elements in iron ore concentrate by wet-fire combined process - Google Patents
Method for extracting valuable elements in iron ore concentrate by wet-fire combined process Download PDFInfo
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- CN114480881A CN114480881A CN202210055777.2A CN202210055777A CN114480881A CN 114480881 A CN114480881 A CN 114480881A CN 202210055777 A CN202210055777 A CN 202210055777A CN 114480881 A CN114480881 A CN 114480881A
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- leaching
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- vanadium
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- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 title claims abstract description 164
- 238000000034 method Methods 0.000 title claims abstract description 95
- 229910052742 iron Inorganic materials 0.000 title claims abstract description 81
- 230000008569 process Effects 0.000 title claims abstract description 50
- 239000012141 concentrate Substances 0.000 title claims abstract description 46
- 239000010936 titanium Substances 0.000 claims abstract description 128
- 238000002386 leaching Methods 0.000 claims abstract description 125
- 229910052719 titanium Inorganic materials 0.000 claims abstract description 118
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 claims abstract description 117
- 229910052720 vanadium Inorganic materials 0.000 claims abstract description 74
- LEONUFNNVUYDNQ-UHFFFAOYSA-N vanadium atom Chemical compound [V] LEONUFNNVUYDNQ-UHFFFAOYSA-N 0.000 claims abstract description 74
- 239000002253 acid Substances 0.000 claims abstract description 60
- 239000008188 pellet Substances 0.000 claims abstract description 59
- 239000000463 material Substances 0.000 claims abstract description 57
- 239000003513 alkali Substances 0.000 claims abstract description 39
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 28
- 238000001354 calcination Methods 0.000 claims abstract description 22
- 239000003245 coal Substances 0.000 claims abstract description 19
- NTHWMYGWWRZVTN-UHFFFAOYSA-N sodium silicate Chemical compound [Na+].[Na+].[O-][Si]([O-])=O NTHWMYGWWRZVTN-UHFFFAOYSA-N 0.000 claims abstract description 16
- 230000009467 reduction Effects 0.000 claims abstract description 15
- 238000010306 acid treatment Methods 0.000 claims abstract description 13
- 238000002844 melting Methods 0.000 claims abstract description 13
- 230000008018 melting Effects 0.000 claims abstract description 13
- 239000004115 Sodium Silicate Substances 0.000 claims abstract description 12
- 229910052911 sodium silicate Inorganic materials 0.000 claims abstract description 12
- 230000001502 supplementing effect Effects 0.000 claims abstract description 4
- 239000000243 solution Substances 0.000 claims description 65
- 239000002893 slag Substances 0.000 claims description 59
- 238000000926 separation method Methods 0.000 claims description 33
- 230000002308 calcification Effects 0.000 claims description 31
- 239000007788 liquid Substances 0.000 claims description 28
- 239000007787 solid Substances 0.000 claims description 27
- 235000021110 pickles Nutrition 0.000 claims description 25
- 239000003795 chemical substances by application Substances 0.000 claims description 20
- 239000003638 chemical reducing agent Substances 0.000 claims description 19
- 238000000605 extraction Methods 0.000 claims description 17
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 16
- 238000006243 chemical reaction Methods 0.000 claims description 16
- 230000001590 oxidative effect Effects 0.000 claims description 13
- 238000007885 magnetic separation Methods 0.000 claims description 12
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical group Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 10
- 239000007800 oxidant agent Substances 0.000 claims description 10
- 238000012216 screening Methods 0.000 claims description 10
- 239000000571 coke Substances 0.000 claims description 8
- 238000002156 mixing Methods 0.000 claims description 7
- 229920006395 saturated elastomer Polymers 0.000 claims description 7
- 239000007790 solid phase Substances 0.000 claims description 7
- VTYYLEPIZMXCLO-UHFFFAOYSA-L Calcium carbonate Chemical compound [Ca+2].[O-]C([O-])=O VTYYLEPIZMXCLO-UHFFFAOYSA-L 0.000 claims description 6
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims description 6
- RHZUVFJBSILHOK-UHFFFAOYSA-N anthracen-1-ylmethanolate Chemical group C1=CC=C2C=C3C(C[O-])=CC=CC3=CC2=C1 RHZUVFJBSILHOK-UHFFFAOYSA-N 0.000 claims description 6
- 239000003830 anthracite Substances 0.000 claims description 6
- 239000011230 binding agent Substances 0.000 claims description 6
- 238000001465 metallisation Methods 0.000 claims description 6
- 229910017604 nitric acid Inorganic materials 0.000 claims description 6
- 238000010992 reflux Methods 0.000 claims description 6
- 238000010790 dilution Methods 0.000 claims description 5
- 239000012895 dilution Substances 0.000 claims description 5
- 238000005453 pelletization Methods 0.000 claims description 5
- 229910052925 anhydrite Inorganic materials 0.000 claims description 4
- 229910000019 calcium carbonate Inorganic materials 0.000 claims description 3
- 239000000920 calcium hydroxide Substances 0.000 claims description 3
- 229910001861 calcium hydroxide Inorganic materials 0.000 claims description 3
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical group [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 claims description 3
- 238000003837 high-temperature calcination Methods 0.000 claims description 3
- 239000000203 mixture Substances 0.000 claims description 3
- 239000000126 substance Substances 0.000 claims description 3
- 238000000227 grinding Methods 0.000 claims description 2
- 239000000155 melt Substances 0.000 claims description 2
- 238000012544 monitoring process Methods 0.000 claims 1
- 239000010413 mother solution Substances 0.000 claims 1
- XUIMIQQOPSSXEZ-UHFFFAOYSA-N Silicon Chemical compound [Si] XUIMIQQOPSSXEZ-UHFFFAOYSA-N 0.000 abstract description 21
- 239000010703 silicon Substances 0.000 abstract description 21
- 229910052710 silicon Inorganic materials 0.000 abstract description 21
- 238000011084 recovery Methods 0.000 abstract description 13
- 239000012535 impurity Substances 0.000 abstract description 12
- 239000000047 product Substances 0.000 description 18
- SZVJSHCCFOBDDC-UHFFFAOYSA-N iron(II,III) oxide Inorganic materials O=[Fe]O[Fe]O[Fe]=O SZVJSHCCFOBDDC-UHFFFAOYSA-N 0.000 description 17
- GWEVSGVZZGPLCZ-UHFFFAOYSA-N Titan oxide Chemical compound O=[Ti]=O GWEVSGVZZGPLCZ-UHFFFAOYSA-N 0.000 description 14
- OKTJSMMVPCPJKN-UHFFFAOYSA-N Carbon Chemical compound [C] OKTJSMMVPCPJKN-UHFFFAOYSA-N 0.000 description 13
- 238000003723 Smelting Methods 0.000 description 12
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 description 12
- GFNGCDBZVSLSFT-UHFFFAOYSA-N titanium vanadium Chemical compound [Ti].[V] GFNGCDBZVSLSFT-UHFFFAOYSA-N 0.000 description 12
- 239000012071 phase Substances 0.000 description 9
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 8
- 239000002994 raw material Substances 0.000 description 7
- 238000005406 washing Methods 0.000 description 7
- 239000012670 alkaline solution Substances 0.000 description 6
- 239000011575 calcium Substances 0.000 description 6
- 239000012452 mother liquor Substances 0.000 description 6
- 239000006227 byproduct Substances 0.000 description 5
- 229910052799 carbon Inorganic materials 0.000 description 5
- 230000004048 modification Effects 0.000 description 5
- 238000012986 modification Methods 0.000 description 5
- 238000004064 recycling Methods 0.000 description 5
- OYPRJOBELJOOCE-UHFFFAOYSA-N Calcium Chemical compound [Ca] OYPRJOBELJOOCE-UHFFFAOYSA-N 0.000 description 4
- 229910000831 Steel Inorganic materials 0.000 description 4
- 229910052791 calcium Inorganic materials 0.000 description 4
- 238000001035 drying Methods 0.000 description 4
- 239000008235 industrial water Substances 0.000 description 4
- 239000007791 liquid phase Substances 0.000 description 4
- MKTRXTLKNXLULX-UHFFFAOYSA-P pentacalcium;dioxido(oxo)silane;hydron;tetrahydrate Chemical compound [H+].[H+].O.O.O.O.[Ca+2].[Ca+2].[Ca+2].[Ca+2].[Ca+2].[O-][Si]([O-])=O.[O-][Si]([O-])=O.[O-][Si]([O-])=O.[O-][Si]([O-])=O.[O-][Si]([O-])=O.[O-][Si]([O-])=O MKTRXTLKNXLULX-UHFFFAOYSA-P 0.000 description 4
- 235000019353 potassium silicate Nutrition 0.000 description 4
- 239000000377 silicon dioxide Substances 0.000 description 4
- 235000012239 silicon dioxide Nutrition 0.000 description 4
- 239000002910 solid waste Substances 0.000 description 4
- 239000010959 steel Substances 0.000 description 4
- 239000010456 wollastonite Substances 0.000 description 4
- 229910052882 wollastonite Inorganic materials 0.000 description 4
- 230000000052 comparative effect Effects 0.000 description 3
- 238000011161 development Methods 0.000 description 3
- 230000000694 effects Effects 0.000 description 3
- 238000005265 energy consumption Methods 0.000 description 3
- 239000000446 fuel Substances 0.000 description 3
- 230000007935 neutral effect Effects 0.000 description 3
- 238000002360 preparation method Methods 0.000 description 3
- 238000003756 stirring Methods 0.000 description 3
- 239000013589 supplement Substances 0.000 description 3
- 239000002699 waste material Substances 0.000 description 3
- 239000004484 Briquette Substances 0.000 description 2
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 description 2
- FYYHWMGAXLPEAU-UHFFFAOYSA-N Magnesium Chemical compound [Mg] FYYHWMGAXLPEAU-UHFFFAOYSA-N 0.000 description 2
- UGACIEPFGXRWCH-UHFFFAOYSA-N [Si].[Ti] Chemical compound [Si].[Ti] UGACIEPFGXRWCH-UHFFFAOYSA-N 0.000 description 2
- 229910052782 aluminium Inorganic materials 0.000 description 2
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 description 2
- 238000004458 analytical method Methods 0.000 description 2
- 230000007547 defect Effects 0.000 description 2
- GNTDGMZSJNCJKK-UHFFFAOYSA-N divanadium pentaoxide Chemical compound O=[V](=O)O[V](=O)=O GNTDGMZSJNCJKK-UHFFFAOYSA-N 0.000 description 2
- 238000005516 engineering process Methods 0.000 description 2
- 238000003912 environmental pollution Methods 0.000 description 2
- 238000007654 immersion Methods 0.000 description 2
- 239000011777 magnesium Substances 0.000 description 2
- 229910052749 magnesium Inorganic materials 0.000 description 2
- WPBNNNQJVZRUHP-UHFFFAOYSA-L manganese(2+);methyl n-[[2-(methoxycarbonylcarbamothioylamino)phenyl]carbamothioyl]carbamate;n-[2-(sulfidocarbothioylamino)ethyl]carbamodithioate Chemical compound [Mn+2].[S-]C(=S)NCCNC([S-])=S.COC(=O)NC(=S)NC1=CC=CC=C1NC(=S)NC(=O)OC WPBNNNQJVZRUHP-UHFFFAOYSA-L 0.000 description 2
- 238000012545 processing Methods 0.000 description 2
- 159000000000 sodium salts Chemical class 0.000 description 2
- 239000004408 titanium dioxide Substances 0.000 description 2
- XHCLAFWTIXFWPH-UHFFFAOYSA-N [O-2].[O-2].[O-2].[O-2].[O-2].[V+5].[V+5] Chemical class [O-2].[O-2].[O-2].[O-2].[O-2].[V+5].[V+5] XHCLAFWTIXFWPH-UHFFFAOYSA-N 0.000 description 1
- 230000002411 adverse Effects 0.000 description 1
- 230000004075 alteration Effects 0.000 description 1
- DNWNZRZGKVWORZ-UHFFFAOYSA-N calcium oxido(dioxo)vanadium Chemical compound [Ca+2].[O-][V](=O)=O.[O-][V](=O)=O DNWNZRZGKVWORZ-UHFFFAOYSA-N 0.000 description 1
- 159000000007 calcium salts Chemical class 0.000 description 1
- 238000001816 cooling Methods 0.000 description 1
- 238000007405 data analysis Methods 0.000 description 1
- 238000006477 desulfuration reaction Methods 0.000 description 1
- 230000023556 desulfurization Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000007865 diluting Methods 0.000 description 1
- 239000011521 glass Substances 0.000 description 1
- 230000007062 hydrolysis Effects 0.000 description 1
- 238000006460 hydrolysis reaction Methods 0.000 description 1
- 229910052500 inorganic mineral Inorganic materials 0.000 description 1
- UQSXHKLRYXJYBZ-UHFFFAOYSA-N iron oxide Inorganic materials [Fe]=O UQSXHKLRYXJYBZ-UHFFFAOYSA-N 0.000 description 1
- 239000000696 magnetic material Substances 0.000 description 1
- 230000005389 magnetism Effects 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 229910052751 metal Inorganic materials 0.000 description 1
- 239000002184 metal Substances 0.000 description 1
- 239000011707 mineral Substances 0.000 description 1
- WWTORYHTBNJMMT-UHFFFAOYSA-N potassium sodium oxygen(2-) Chemical compound [K+].[O-2].[Na+] WWTORYHTBNJMMT-UHFFFAOYSA-N 0.000 description 1
- 239000000843 powder Substances 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 150000003839 salts Chemical class 0.000 description 1
- 238000002791 soaking Methods 0.000 description 1
- 238000009628 steelmaking Methods 0.000 description 1
- 150000003681 vanadium Chemical class 0.000 description 1
- 229910001935 vanadium oxide Inorganic materials 0.000 description 1
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B34/00—Obtaining refractory metals
- C22B34/10—Obtaining titanium, zirconium or hafnium
- C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
- C22B34/1236—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
- C22B34/124—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors
- C22B34/125—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors containing a sulfur ion as active agent
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
- C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—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
- C22B34/10—Obtaining titanium, zirconium or hafnium
- C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
- C22B34/1218—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by dry processes
-
- 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
- C22B34/10—Obtaining titanium, zirconium or hafnium
- C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
- C22B34/1236—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
- C22B34/124—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors
- C22B34/1245—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors containing a halogen ion as active agent
-
- 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
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- 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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- Mechanical Engineering (AREA)
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- Inorganic Chemistry (AREA)
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Abstract
The invention discloses a method for extracting valuable elements in iron ore concentrate by a wet-fire combined process, which comprises the following steps: s1, extracting vanadium by a wet method; s2, pellet calcination; s3, reducing by using a coal-based shaft furnace; s4, sorting materials; s5, electric furnace melting and separating; s6, dilute acid treatment; s7, supplementing water and leaching; and S8, alkali treatment. The method of firstly separating vanadium and then separating iron is adopted, so that vanadium is firstly separated, the influence of vanadium on the subsequent process is reduced, and the recovery rate of iron, silicon and titanium is improved. Then, coal-based shaft furnace reduction and electric furnace melting are adopted to replace the existing blast furnace method, and finally, a dilute acid treatment and water supplementing leaching method is adopted to replace the existing conventional acid leaching method, so that the impurity removal rate reaches over 95 percent, the recovery rate of titanium is improved, and finally, titanium and silicon are separated through alkali treatment to obtain a high-purity sodium silicate solution and a titanium-rich material.
Description
Technical Field
The invention belongs to the technical field of titanium recycling, and particularly relates to a method for extracting valuable elements in iron ore concentrate by a wet-fire combined process.
Background
Valuable resources in the vanadium titano-magnetite are resources such as iron, vanadium, titanium and the like, and the blast furnace method is mainly used in the technical field of processing vanadium titano-magnetite resources at present, but the blast furnace method has high energy consumption, great environmental pollution and low resource utilization rate and is inevitably replaced by non-blast furnace smelting. A reasonable way of smelting vanadium-titanium magnetite by adopting a non-blast furnace iron-making technology is explored to further realize comprehensive utilization of iron, vanadium and titanium resources, and the new process can be divided into a process of firstly extracting vanadium and then iron, a process of firstly extracting iron and then vanadium and a process of simultaneously extracting iron, vanadium and titanium according to different element extraction sequences.
The vanadium-iron-first flow process is to calcine the vanadium-titanium magnetite concentrate by sodium (potassium) oxide to convert vanadium into soluble vanadium salt for leaching. And reducing the residual balls in a rotary kiln and melting and separating in an electric furnace to obtain molten steel and titanium slag, so that iron, vanadium and titanium are recycled. The process has the advantages that the comprehensive utilization degree of iron, vanadium and titanium is high, but the process flow is still long, and the large process problem of the pellets containing sodium salt in the iron-making process cannot be solved.
The process comprises the steps of firstly carrying out pre-reduction on vanadium-titanium magnetite concentrate, carrying out melt separation or magnetic separation to obtain a metal iron material, further steelmaking, mainly enriching vanadium and titanium into slag, and then extracting by the traditional vanadium-titanium extraction process. The main process comprises the procedures of shaft furnace-electric furnace, rotary kiln-electric furnace, rotary hearth furnace-electric furnace and the like. The process flow is still longer, the melting separation cost is high, and the resource utilization difficulty of vanadium and titanium in the slag is still larger.
According to the vanadium titano-magnetite processing technology, most of the existing methods can only effectively recover one or two elements, the resource utilization rate is low, most of the processes need high-temperature reduction treatment on iron ore concentrate, and the cost and the energy consumption are high, so that the research on the high-efficiency clean extraction method of valuable elements of the iron ore concentrate has very important significance for the development of iron, titanium and vanadium industries.
The blast furnace method is the most common method for treating vanadium titano-magnetite resources and the most mature method for treating the vanadium titano-magnetite resources in China at present. The method comprises the steps of firstly, feeding sintered ore and pellet ore and solid reducing agents (such as coke and coal powder) into a blast furnace for smelting, then selectively reducing iron and vanadium oxides in the ore through the blast furnace smelting to generate vanadium-containing molten iron, and enriching titanium in blast furnace slag. The vanadium-containing molten iron is subjected to pre-desulfurization and then blown by a converter, vanadium is oxidized to form vanadium slag, the molten iron is changed into semisteel, the vanadium slag is subjected to a wet process to obtain a qualified vanadium product, and the semisteel is further processed to obtain a qualified steel product. The titanium slag is difficult to be effectively recycled due to the characteristics of low titanium dioxide content, complex mineral phase, more glass phase and the like. The method has the advantages that most of iron in the vanadium-titanium magnetite concentrate can be recycled, and the method has the defects of high production cost and energy consumption, high dependence on coke, incapability of effectively recycling high-titanium slag, environmental pollution and waste of vanadium-titanium resources.
The Pan steel group utilizes the existing process flow to process vanadium-titanium magnetite resources, and only can obtain indexes that the recovery rates of iron, vanadium and titanium elements are respectively 60%, 39% and 10% from raw ores to steel billets, flaky vanadium pentoxide and titanium dioxide products. From the perspective of efficient and green development of resources, the blast furnace method for treating vanadium titano-magnetite does not meet the basic national conditions of China at the present stage, and is not suitable for being used as a development direction for comprehensive utilization of vanadium titano-magnetite.
The invention aims to fully utilize resources such as iron, vanadium, titanium and the like in the vanadium-titanium magnetite concentrate, reduce the solid waste yield, maximize the value of by-products and minimize the harm to the environment.
Disclosure of Invention
The invention aims to provide a method for extracting valuable elements in iron ore concentrate by a wet-fire combined process to overcome the defects of the prior art.
The purpose of the invention is realized by the following technical scheme:
a method for extracting valuable elements in iron ore concentrate by a wet-fire combined process comprises the following steps:
s1, wet vanadium extraction: mixing iron ore concentrate with a calcification agent and a binder, pelletizing, carrying out high-temperature calcification roasting on the pellets at 1100-1300 ℃, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and leached pellets; the using amount of the calcification agent is 0.8-3% of the mass of the iron ore concentrate; the low-concentration acid solution H+The concentration is 0.3-1.5 mol/L;
s2, pellet calcination: calcining the dipped pellets at a high temperature of 1200-1300 ℃;
s3, coal-based shaft furnace reduction: mixing the pellets with a solid reducing agent, and reducing the mixture in a coal-based shaft furnace reactor at 800-1300 ℃ for 12-25 h to obtain metallized pellets with the metallization rate of more than or equal to 90%;
s4, material sorting: screening the material reduced in the step S3, carrying out magnetic separation on the screened material, and collecting the non-magnetic reducing agent;
s5, electric furnace melting separation: feeding the material sieved in the step S4 into an electric furnace, and carrying out melt separation at 1450-1700 ℃, wherein the melt separation time is 50-90 min, so that iron and slag are separated, and titanium-containing slag and molten iron are obtained;
s6, dilute acid treatment: placing the titanium-containing slag into acid leaching solution, performing reflux reaction for 3-8 h at 90-120 ℃, adding an oxidant after the reaction for 1-2 h to ensure that Ti in the solution is3+Oxidized to Ti4+(ii) a The mass percentage concentration of acid in the pickle liquor is 15-50%, and the liquid-solid ratio is 3-10: 1;
s7, water replenishing and leaching: adding water for dilution, wherein the water addition amount is 1-1.5 times of the volume of the pickle liquor, continuously reacting for 1-2 hours at a reaction temperature not lower than the diluted acid treatment in the step S6, and performing solid-liquid separation after leaching to obtain leached liquor and leached residues;
s8, alkali treatment: and (3) placing the leached residues into an alkali leaching solution, and leaching for 3-6 hours at 40-80 ℃, wherein the mass fraction of alkali in the alkali leaching solution is 10-25%.
Preferably, the calcification agent in step S1 is CaSO4、CaCO3、Ca(OH)2One or a combination of two or more of (1).
Preferably, in the step S1, the calcification roasting time is 1.5-3 h, the acid leaching temperature is 10-50 ℃, and the leaching time is more than or equal to 15 d.
Preferably, the high-temperature calcination time in the step S2 is 0.5 to 1.5 hours.
Preferably, the solid reducing agent in step S3 is anthracite or coke.
Preferably, in the step S4, the screening mesh number is 6-10 meshes, and the magnetic separation intensity is 1800-2300 Gs.
Preferably, the titanium-containing slag is firstly ground to-200 meshes of more than or equal to 80 percent before acid leaching in the step S6.
Preferably, the pickle liquor in step S6 is dilute hydrochloric acid, dilute sulfuric acid or dilute nitric acid, the mass percentage concentration of the pickle liquor is 15-25% when the pickle liquor is dilute hydrochloric acid, 20-50% when the pickle liquor is dilute sulfuric acid, and 20-40% when the pickle liquor is dilute nitric acid.
Preferably, the concentration of alkali is monitored in the alkali leaching process in step S8, and when the concentration of alkali is lower than 15%, a certain amount of saturated alkali liquor is added to ensure that the concentration of alkali is greater than or equal to 15%.
Preferably, after the alkali treatment in step S8, a titanium-containing solid phase and a sodium silicate mother liquor are obtained through solid-liquid separation, and the titanium-containing solid phase is calcined at 850-1000 ℃ to obtain a titanium-rich material.
Compared with the prior art, the invention has the following advantages:
(1) different from the traditional extraction process of vanadium in iron ore concentrate, the wet method vanadium extraction is carried out before iron making, the process cost is low, the recovery rate of vanadium can reach more than 65 percent, and the non-blast furnace reduction smelting is adopted, so that the energy utilization rate is high.
(2) The value of byproducts can be fully utilized, for example, materials reduced by the coal-based shaft furnace, and nonmagnetic reducing agents after magnetic separation can be used as active carbon, so that the value is improved. And the alkaline sodium silicate solution obtained after alkaline leaching can be used for preparing silicon-containing materials, the utilization rate of waste byproducts is high, and the generation amount of solid waste is zero.
(3) The utilization rate of titanium in the titanium-containing slag generated in the smelting process is high, the recovery rate of titanium is more than or equal to 94 percent, and the grade of titanium of the obtained titanium-rich product is more than or equal to 80 percent.
(4) Fully realizes the full utilization of iron, titanium and vanadium resources in the iron ore concentrate.
Drawings
FIG. 1 is a process flow diagram provided by the present invention.
Detailed Description
A method for extracting valuable elements in iron ore concentrate by a wet-fire combined process is shown in figure 1 and comprises the following steps:
s1, wet vanadium extraction: mixing iron ore concentrate with calcification agent and binder, pelletizingCarrying out high-temperature calcification roasting at 1100-1300 ℃, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and leached pellets, wherein the vanadium-containing solution can be used for subsequent vanadium extraction, and the leached pellets are dried and then used for subsequent smelting; the using amount of the calcification agent is 0.8-3% of the mass of the iron ore concentrate; low concentration acid solution H+The concentration is 0.3-1.5 mol/L;
the calcium salt is adopted for calcified roasting, so that the adverse effect of roasting the sodium salt on pellet iron making can be solved;
s2, pellet calcination: calcining the pellets after soaking at the high temperature of 1200-1300 ℃ to improve the strength of the pellets, wherein the compressive strength is more than or equal to 1200N/pellet;
s3, coal-based shaft furnace reduction: mixing the pellets with a solid reducing agent, and reducing the mixture in a coal-based shaft furnace reactor at 800-1300 ℃ for 12-25 h to obtain metallized pellets with the metallization rate of more than or equal to 90%;
s4, material sorting: screening the material reduced in the step S3, magnetically separating the screened material, and collecting the nonmagnetic reducing agent which can be used for preparing activated carbon;
s5, electric furnace melting separation: feeding the material sieved in the step S4 into an electric furnace, and carrying out melt separation at 1450-1700 ℃ for 50-90 min to separate iron from slag to obtain TiO2The content is more than or equal to 25 percent and contains titanium slag and molten iron;
s6, dilute acid treatment: placing the titanium-containing slag into acid leaching solution, performing reflux reaction at 90-120 ℃ for 3-8 h to remove most of acid-soluble impurities (such as calcium, magnesium, aluminum, manganese, iron and the like), and adding an oxidant after the reaction is performed for 1-2 h to ensure that Ti in the solution is Ti3+Oxidized to Ti4+(ii) a The acid mass percentage concentration of the pickle liquor is 15-50%, and the liquid-solid ratio is 3-10: 1; the oxidant can be selected from one or more of ferric salt and hydrogen peroxide, and the addition amount of the oxidant is 2-10% of the total mass of the titanium-containing slag;
s7, water replenishing and leaching: adding water for dilution, wherein the added water amount is 1-1.5 times of the volume of the pickle liquor, continuously reacting for 1-2 hours at a reaction temperature not lower than the diluted acid treatment in the step S6, and performing solid-liquid separation after leaching to obtain leached liquor and leached residues; titanium is enriched in a slag phase through dilute acid treatment and water supplementing leaching, and the yield of the titanium is more than or equal to 94 percent;
the impurities such as calcium, magnesium, aluminum, manganese, iron and the like in the titanium-containing slag can be effectively removed through dilute acid leaching, the reaction activity of the impurities and the dilute acid is high, the removal rate is over 95 percent, but silicon cannot be removed through acid leaching and is still in a slag phase, so that the slag after leaching mainly comprises titanium and silicon. Because the titanium-containing slag contains a certain amount of low-valence titanium Ti3+，Ti3+Is more reactive than Ti4+High, dilute acid leaching Ti3+Will go into solution causing loss of titanium, if an oxidant is added during leaching, will add Ti3+Is oxidized to Ti4+Adding a certain amount of water to dilute the leaching solution, and continuing to react for a period of time to obtain Ti4+Hydrolysis and return to the slag phase (if only oxidant is added, the oxidized Ti can not be hydrolyzed into the slag phase), thus the yield of the titanium can reach more than 94% in the dilute acid leaching stage.
S8, alkali treatment: and (3) placing the residues after leaching into an alkaline leaching solution, wherein the alkaline leaching solution can be prepared from NaOH, KOH and the like, and is leached for 3-6 hours at the temperature of 40-80 ℃, and the mass fraction of alkali in the alkaline leaching solution is 10-25%.
Dissolving silicon in the leached slag through alkali treatment; and after the alkaline leaching is finished, performing solid-liquid separation, wherein the titanium-containing material is left in the solid, and the silicon enters the liquid phase, so that the separation of the titanium and the silicon is realized. Washing the solid with water to neutrality, and drying. The mother liquor after alkaline leaching is alkaline sodium silicate solution, has low impurity content, can be used as raw materials of industrial water glass, and can also be used for preparing wollastonite, tobermorite, silicon dioxide and other materials. Calcining the titanium-containing solid phase at 850-1000 ℃, and optimizing the process, such as adding alkali in the alkali leaching process to ensure the concentration of the alkali, so as to obtain a titanium-rich material product containing TiO2≥85％。
According to the invention, firstly, calcification roasting is adopted to convert vanadium in iron ore concentrate into calcium vanadate, and then low-concentration acid leaching is carried out to dissolve vanadium in the solution as far as possible, so as to obtain a high-yield vanadium solution. And then calcining the dipped pellets at high temperature to improve the pressure of the pellets so that the pellets have enough strength to cope with the pressure of the coal-based shaft furnace, reducing iron in the materials into simple substance iron after the reduction of the coal-based shaft furnace, screening and magnetically separating the reduced materials, recovering a reducing agent, and preparing active carbon.
Therefore, the method of firstly separating vanadium and then separating iron is adopted, so that the vanadium is firstly separated, the influence of vanadium on the subsequent process is reduced, and the recovery rate of iron, silicon and titanium is improved. And finally, a diluted acid treatment and water supplement leaching method is adopted to replace the conventional acid leaching method in the prior art, so that the impurity removal rate reaches over 95 percent, the recovery rate of titanium is improved, and finally, the titanium and the silicon are separated through alkali treatment to obtain a high-purity sodium silicate solution and a titanium-rich material. The recovery rate of vanadium is more than or equal to 65%, the recovery rate of iron is more than or equal to 90%, the recovery rate of titanium is more than or equal to 94%, the recovery rate of silicon is more than or equal to 80%, and no solid waste is generated in the process.
Compared with the prior art, the invention has the following advantages:
(1) different from the traditional extraction process of vanadium in iron ore concentrate, the wet method vanadium extraction is carried out before iron making, the process cost is low, the recovery rate of vanadium can reach more than 70 percent, and the non-blast furnace reduction smelting is adopted, so that the energy utilization rate is high.
(2) The value of byproducts can be fully utilized, for example, materials reduced by the coal-based shaft furnace, and nonmagnetic reducing agents after magnetic separation can be used as active carbon, so that the value is improved. And the alkaline sodium silicate solution obtained after alkaline leaching can be used for preparing silicon-containing materials, the utilization rate of waste byproducts is high, and the generation amount of solid waste is zero.
(3) The utilization rate of titanium in the titanium-containing slag generated in the smelting process is high, the recovery rate of titanium is more than or equal to 94 percent, and the grade of titanium of the obtained titanium-rich product is more than or equal to 80 percent.
(4) Fully realizes the full utilization of iron, titanium and vanadium resources in the iron ore concentrate.
Preferably, the calcification agent in step S1 is CaSO4、CaCO3、Ca(OH)2One or a combination of two or more of (1).
Preferably, the calcification roasting time of the step S1 is 1.5-3 h, the acid leaching temperature is 10-50 ℃, and the leaching time is more than or equal to 15 d.
Preferably, the high-temperature calcination time of the step S2 is 0.5-1.5 h.
Preferably, the solid reducing agent in step S3 is anthracite or coke breeze, and can be recovered as activated carbon after being reduced by screening and magnetic separation.
Preferably, in the step S4, the screening mesh number is 6-10 meshes, the magnetic separation strength is 1800-2300 Gs, and the reduced carbon without magnetism can be screened and separated in the magnetic field strength range. The magnetic material is the broken material stripped from the pellet, which can be collected to be pelletized and enter the next step together with other pellets.
Preferably, before acid leaching, the titanium-containing slag is ground in step S6 to a content of-200 meshes of 80% or more, and the acid leaching efficiency can be improved by grinding.
Preferably, the pickle liquor in step S6 is dilute hydrochloric acid, dilute sulfuric acid or dilute nitric acid, the mass percentage concentration of the pickle liquor is 15-25% when the pickle liquor is dilute hydrochloric acid, 20-50% when the pickle liquor is dilute sulfuric acid, and 20-40% when the pickle liquor is dilute nitric acid.
Preferably, the concentration of alkali is monitored in the alkali leaching process in step S8, and when the concentration of alkali is lower than 15%, a certain amount of saturated alkali liquor is added to ensure that the concentration of alkali is greater than or equal to 15%. The alkaline leaching efficiency can be improved by ensuring the concentration of the alkali.
Preferably, after the alkali treatment in step S8, a titanium-containing solid phase and a sodium silicate mother liquor are obtained through solid-liquid separation, and the titanium-containing solid phase is calcined at 850-1000 ℃ to obtain a titanium-rich material.
Example 1
The method for extracting the valuable elements from the vanadium-titanium magnetite concentrate to obtain the molten iron, the vanadium-containing liquid and the titanium-rich material comprises the following specific preparation steps:
A. and (3) wet vanadium extraction: using iron ore concentrate as raw material, CaSO4The calcium agent is a calcification agent, iron ore concentrate, the calcification agent and a binder are mixed according to a certain proportion, and pelletizing is carried out by a pelletizer. Ball with ball-shaped sectionAnd (3) carrying out high-temperature calcification roasting on the pellets, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and leached pellets, wherein the vanadium-containing solution is used for subsequent vanadium extraction, and the leached pellets are dried and then used for subsequent smelting. The addition of calcification agent is 1.5% of iron ore concentrate, the calcification roasting temperature is 1200 deg.C, the time is 2H, the acid leaching solution is dilute sulfuric acid, H+The concentration is 0.9mol/L, the liquid-solid ratio is 3:1, the leaching temperature is 25 ℃, the leaching time is 25 d. The leaching rate of vanadium is determined to be 65%, the obtained vanadium-containing liquid is concentrated, and the concentration of V is 2.5 g/L.
B. Pellet calcination: and drying the pellets after the immersion, and calcining the pellets at 1250 ℃ for 1h to improve the strength of the pellets, wherein the compressive strength is more than or equal to 1250N/pellet.
C. Reducing in a coal-based shaft furnace: the pellets and a solid reducing agent (anthracite or coke briquette) are mixed, mixed furnace burden is added into a coal-based vertical reactor through a feeding system, fuel is combusted in a furnace through a burner to heat the furnace burden, the furnace burden is preheated and reduced in the closed vertical reactor, the reduction temperature is about 1220 ℃, the reduction time is 18 hours, and the metallized pellets with the metallization rate of 92% are obtained.
D. Material sorting: the reduced materials are cooled and sieved (the mesh number is 8), the sieved materials are subjected to magnetic separation (the strength is 2000Gs), and nonmagnetic reducing agents are collected and can be used for preparing activated carbon.
E. Electric furnace melting and separating: feeding the oversize pellets into an electric furnace, melting at 1650 deg.C for 60min to separate iron from slag to obtain TiO2Titanium-containing slag and molten iron with the content of 27 percent.
F. Treating with dilute acid: and (3) placing the ground titanium-containing slag into the pickle liquor, continuously stirring, carrying out reflux reaction for 6 hours, and removing most impurities. The proportion of-200 meshes of the ground blast furnace slag is more than or equal to 83 percent. The acid leaching solution is dilute hydrochloric acid, the concentration is 21%, the liquid-solid ratio is 6:1(mL/g), and the leaching temperature is 110 ℃. Adding a certain amount of oxidant FeCl in the acid leaching process3(accounting for 3 percent of the total amount of the titanium-containing slag), oxidizing the low-valence titanium in the titanium-containing slag, recycling the leached liquid, and preparing the dilute acid leaching liquid by the leached liquid and concentrated acid.
G. Water replenishing and leaching: diluting with water, wherein the water addition amount is 1.2 times of the volume of the pickle liquor, continuously reacting for 1.5h at a temperature (115 ℃) which is not lower than the dilute acid leaching reaction temperature, carrying out solid-liquid separation after leaching to obtain leached liquor and leached slag, and enriching titanium into a slag phase in the dilute acid treatment and water supplement leaching stages, wherein the titanium yield is 95.5%. Washing the leached slag to neutrality and stoving.
H. Alkali treatment: and (2) placing the dried material into an alkaline leaching solution, wherein the alkaline leaching solution is prepared from NaOH, the concentration of the alkaline leaching solution is 16%, the alkaline leaching temperature is 55 ℃, the leaching time is 4.5 hours, and a certain amount of saturated alkaline solution (the mass percentage concentration is 52%) is supplemented in the leaching process, so that the concentration of the alkaline solution is more than or equal to 15%.
I. Separating titanium and silicon: and after the alkaline leaching is finished, performing solid-liquid separation, wherein the titanium-containing material is left in the solid, and the silicon enters the liquid phase, so that the separation of the titanium and the silicon is realized. And washing the solid to be neutral, and calcining to obtain the titanium-rich material. The mother liquor after alkaline leaching is alkaline sodium silicate solution, has low impurity content, can be used as raw materials of industrial water glass, and can also be used for preparing wollastonite, tobermorite, silicon dioxide and other materials. The calcination temperature was 900 ℃. The obtained titanium-rich material product contains TiO2The content was found to be 83.2%.
Example 2
The method for extracting the valuable elements from the vanadium-titanium magnetite concentrate to obtain the molten iron, the vanadium-containing liquid and the titanium-rich material comprises the following specific preparation steps:
A. and (3) wet vanadium extraction: using iron ore concentrate as raw material, Ca (OH)2The calcium agent is a calcification agent, iron ore concentrate, the calcification agent and a binder are mixed according to a certain proportion, and pelletizing is carried out by a pelletizer. And carrying out high-temperature calcification roasting on the pellets, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and the leached pellets, wherein the vanadium-containing solution is used for subsequent vanadium extraction, and the leached pellets are dried and then used for subsequent smelting. The addition of the calcification agent is 2 percent of the iron ore concentrate, the calcification roasting temperature is 1250 ℃, the acid leaching solution is dilute hydrochloric acid, H is 1.5H+The concentration is 0.5mol/L, the liquid-solid ratio is 2.5:1, the leaching temperature is 22 ℃, and the leaching time is 30 d. The leaching rate of vanadium is 68 percent, and the concentration of V in the obtained vanadium-containing liquid is 1.8 g/L.
B. Pellet calcination: and drying the pellets after the leaching, and calcining the pellets at 1250 ℃ for 1h to improve the strength of the pellets, wherein the compressive strength is more than or equal to 1230N/pellet.
C. Reducing in a coal-based shaft furnace: the pellets and a solid reducing agent (anthracite or coke breeze) are mixed, mixed furnace burden is added into a coal-based vertical reactor through a feeding system, fuel is combusted in a furnace through a burner to heat the furnace burden, the furnace burden is preheated and reduced in the closed vertical reactor, the reduction temperature is about 1190 ℃, the reduction time is 16 hours, and the metallized pellets with the metallization rate of 91.2 percent are obtained.
D. Material sorting: the reduced material is cooled and sieved (the mesh number is 10), the sieved material is magnetically separated (the magnetic separation strength is 2200Gs), and a nonmagnetic reducing agent is collected and can be used for preparing the active carbon.
E. Electric furnace melting separation: and (3) feeding the oversize pellets into an electric furnace, and carrying out melting separation at 1680 ℃ for 90min to separate iron from slag. To obtain TiO2The titanium-containing slag with the content of 30 percent and molten iron.
F. Treating with dilute acid: and (3) placing the ground titanium-containing slag into the pickle liquor, continuously stirring, carrying out reflux reaction for 5 hours, and removing most impurities. The proportion of-200 meshes of the ground blast furnace slag is more than or equal to 81 percent. The acid leaching solution is dilute hydrochloric acid, the concentration is 19%, the liquid-solid ratio is 6:1(mL/g), and the leaching temperature is 100 ℃. A certain amount of oxidant FeCl can be added in the acid leaching process3(4% of the total amount of the titanium-containing slag), oxidizing the low-valence titanium in the titanium-containing slag, recycling the leached liquid, and preparing the leached liquid into dilute acid leaching liquid with concentrated acid.
G. Water replenishing and leaching: adding water for dilution, wherein the water addition amount is 1.2 times of the volume of the pickle liquor, continuously reacting for 2 hours at a reaction temperature (110 ℃) which is not lower than the dilute acid leaching in the step, carrying out solid-liquid separation after the leaching is finished to obtain leached liquor and leached slag, and enriching titanium into a slag phase in the stages of dilute acid treatment and water supplement leaching, wherein the yield of the titanium is 94.8%. Washing the leached slag to neutrality and stoving.
H. Alkali treatment: placing the dried material in alkaline leaching solution, wherein the alkaline leaching solution can be prepared from NaOH, the concentration of the alkaline leaching solution is 15%, the alkaline leaching temperature is 60 ℃, the leaching time is 4h, and a certain amount of saturated alkaline solution (the concentration is 52%) is supplemented in the leaching process to ensure that the concentration of the alkaline solution is more than or equal to 15%.
I. And (3) titanium-silicon separation: after the alkaline leaching is finishedAnd (4) carrying out solid-liquid separation, wherein the titanium-containing material is left in the solid, and the silicon enters the liquid phase, so that the separation of the titanium and the silicon is realized. And washing the solid to be neutral, and calcining to obtain the titanium-rich material. The mother liquor after alkaline leaching is alkaline sodium silicate solution, has low impurity content, can be used as raw materials of industrial water glass, and can also be used for preparing wollastonite, tobermorite, silicon dioxide and other materials. The calcination temperature was 950 ℃. The obtained titanium-rich material product contains TiO2The content was 87.3%.
Example 3
The method for extracting the valuable elements from the vanadium-titanium magnetite concentrate to obtain the molten iron, the vanadium-containing liquid and the titanium-rich material comprises the following specific preparation steps:
A. and (3) wet vanadium extraction: using vanadium-titanium magnetite concentrate as raw material, Ca (OH)2And the like as a calcification agent, mixing the iron ore concentrate, the calcification agent and a binder in a certain proportion, and pelletizing by using a pelletizer. And carrying out high-temperature calcification roasting on the pellets, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and the leached pellets, wherein the vanadium-containing solution is used for subsequent vanadium extraction, and the leached pellets are dried and then used for subsequent smelting. The addition amount of the calcification agent is 2% of the iron ore concentrate, the calcification roasting temperature is 1240 ℃, the roasting temperature is 2h, the acid leaching solution is dilute sulfuric acid, and the liquid-solid ratio is 2: 1, H+The concentration is 1.1mol/L, and the leaching time is 27 days. The leaching rate of the vanadium is 70 percent, and the concentration of V in the obtained vanadium-containing liquid is 2 g/L.
B. Pellet calcination: drying the pellets after immersion, and calcining the pellets at 1200 ℃ for 1h at high temperature to improve the pellet strength, wherein the compressive strength is more than or equal to 1200N/pellet.
C. Reducing by using a coal-based shaft furnace: the pellets and a solid reducing agent (anthracite or coke briquette) are mixed, mixed furnace burden is added into a coal-based vertical reactor through a feeding system, fuel is combusted in a furnace through a burner to heat the furnace burden, the furnace burden is preheated and reduced in the closed vertical reactor, the reduction temperature is 1200 ℃, the reduction time is 17 hours, and the metallized pellets with the metallization rate of 90.8 percent are obtained.
D. Material sorting: and cooling the reduced material, screening (the screening mesh number is 8), magnetically separating the screened material (the magnetic separation strength is 2200Gs), and collecting a nonmagnetic reducing agent which can be used for preparing the activated carbon.
E. Electric furnace melting separation: and (3) feeding the oversize pellets into an electric furnace, and carrying out melting separation at 1600 ℃ for 80min to separate iron from slag. To obtain TiO2The content of the titanium-containing slag is 33 percent.
F. Treating with dilute acid: and (3) placing the ground titanium-containing slag into the pickle liquor, continuously stirring, carrying out reflux reaction for 5 hours, and removing most impurities. The proportion of-200 meshes of the ground blast furnace slag is more than or equal to 81 percent. The acid leaching solution is dilute hydrochloric acid, the concentration is 19%, the liquid-solid ratio is 6:1(mL/g), and the leaching temperature is 100 ℃. A certain amount of oxidant FeCl can be added in the acid leaching process3(accounting for 7 percent of the total amount of the titanium-containing slag), oxidizing the low-valence titanium in the titanium-containing slag, recycling the leached liquid, and preparing the leached liquid into dilute acid leaching liquid with concentrated acid.
G. Water replenishing and leaching: adding water for dilution, wherein the water addition amount is 1.2 times of the volume of the acid leaching solution, continuously reacting for 1h at a temperature (120 ℃) which is not lower than the reaction temperature of the dilute acid leaching, and carrying out solid-liquid separation after the leaching is finished to obtain a leached solution and leached slag, wherein titanium is enriched in a slag phase, and the yield of the titanium is 94.2%. Washing the leached slag to neutrality and stoving.
H. Alkali treatment: and (3) placing the dried material into an alkaline leaching solution, wherein the alkaline leaching solution can be prepared from NaOH, the concentration of the alkaline leaching solution is 15%, the alkaline leaching temperature is 60 ℃, the leaching time is 4 hours, and a certain amount of saturated alkaline solution (the mass percentage concentration is 52%) is supplemented in the leaching process to ensure that the concentration of the alkaline solution is more than or equal to 15%.
I. And (3) titanium-silicon separation: and after the alkaline leaching is finished, performing solid-liquid separation, wherein the titanium-containing material is left in the solid, and the silicon enters the liquid phase, so that the separation of the titanium and the silicon is realized. And washing the solid to be neutral, and calcining to obtain the titanium-rich material. The mother liquor after alkaline leaching is alkaline sodium silicate solution, has low impurity content, can be used as raw materials of industrial water glass, and can also be used for preparing wollastonite, tobermorite, silicon dioxide and other materials. The calcination temperature was 950 ℃. The obtained titanium-rich material product contains TiO2The content was 85.0%.
Comparative example 1
Compared with the example 2, after the dilute acid treatment, the water replenishing leaching is not carried out, other steps and parameters are consistent, the oxidized titanium is not hydrolyzed, and part of the titanium dissolved in the dilute acid is not effectively recovered. Guide tubeThe yield of titanium after dilute acid leaching is 75.6 percent and is obviously lower than that of a sample after water replenishing leaching, and the finally obtained titanium-rich material product contains TiO274.5%, the titanium grade of this product is also lower than that of the titanium product of example 2.
Comparative example 2
Compared with the example 2, in the alkali treatment stage, saturated alkali liquor is not added in the leaching process, the concentration of alkali in the leaching solution is lower than 10% after leaching for 2 hours, and other steps and parameters are consistent. The titanium product obtained by the method has poor silicon removal effect and high silicon content, so that the titanium content of the titanium product is only 72.6%, and the titanium grade of the product is obviously lower than that of the titanium product in the embodiment 2.
The index analysis of the titanium products obtained in the specific examples 1 to 3 and comparative examples 1 to 2 is shown in tables 1 to 2,
TABLE 1 titanium-rich product index analysis
Table 2 sodium silicate solution ICP data analysis
While preferred embodiments of the present invention have been described, additional variations and modifications in those embodiments may occur to those skilled in the art once they learn of the basic inventive concepts. Therefore, it is intended that the appended claims be interpreted as including preferred embodiments and all such alterations and modifications as fall within the scope of the invention. It will be apparent to those skilled in the art that various changes and modifications may be made in the present invention without departing from the spirit and scope of the invention. Thus, if such modifications and variations of the present invention fall within the scope of the claims of the present invention and their equivalents, the present invention is also intended to include such modifications and variations.
Claims (10)
1. A method for extracting valuable elements in iron ore concentrate by a wet-fire combined process is characterized by comprising the following steps:
s1, wet vanadium extraction: mixing iron ore concentrate with a calcification agent and a binder, pelletizing, carrying out high-temperature calcification roasting on the pellets at 1100-1300 ℃, and carrying out acid leaching by using a low-concentration acid solution to obtain a vanadium-containing solution and leached pellets; the using amount of the calcification agent is 0.8-3% of the mass of the iron ore concentrate; the low-concentration acid solution H+The concentration is 0.3-1.5 mol/L;
s2, pellet calcination: calcining the dipped pellets at a high temperature of 1200-1300 ℃;
s3, coal-based shaft furnace reduction: mixing the pellets with a solid reducing agent, and reducing the mixture in a coal-based shaft furnace reactor at 800-1300 ℃ for 12-25 h to obtain metallized pellets with the metallization rate of more than or equal to 90%;
s4, material sorting: screening the material reduced in the step S3, carrying out magnetic separation on the screened material, and collecting the non-magnetic reducing agent;
s5, electric furnace melting separation: feeding the material sieved in the step S4 into an electric furnace, and carrying out melt separation at 1450-1700 ℃, wherein the melt separation time is 50-90 min, so that iron and slag are separated, and titanium-containing slag and molten iron are obtained;
s6, dilute acid treatment: placing the titanium-containing slag into acid leaching solution, performing reflux reaction for 3-8 h at 90-120 ℃, adding an oxidant after the reaction for 1-2 h to ensure that Ti in the solution is3+Oxidized to Ti4+(ii) a The mass percentage concentration of acid in the pickle liquor is 15-50%, and the liquid-solid ratio is 3-10: 1;
s7, water replenishing and leaching: adding water for dilution, wherein the water addition amount is 1-1.5 times of the volume of the pickle liquor, continuously reacting for 1-2 hours at a reaction temperature not lower than the diluted acid treatment in the step S6, and performing solid-liquid separation after leaching to obtain leached liquor and leached residues;
s8, alkali treatment: and (3) placing the leached residues into an alkali leaching solution, and leaching for 3-6 hours at 40-80 ℃, wherein the mass fraction of alkali in the alkali leaching solution is 10-25%.
2. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
calcification in step S1The agent is selected from CaSO4、CaCO3、Ca(OH)2One or a combination of two or more of (1).
3. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and S1, the calcification roasting time is 1.5-3 h, the acid leaching temperature is 10-50 ℃, and the leaching time is more than or equal to 15 d.
4. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and step S2, the high-temperature calcination time is 0.5-1.5 h.
5. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
in step S3, the solid reducing agent is anthracite or coke.
6. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and S4, the screening mesh number is 6-10 meshes, and the magnetic separation strength is 1800-2300 Gs.
7. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and S6, grinding the titanium-containing slag before acid leaching until the content of the ground titanium-containing slag is more than or equal to 80 percent in a-200-mesh ratio.
8. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and step S6, the pickle liquor is dilute hydrochloric acid, dilute sulfuric acid or dilute nitric acid, when the pickle liquor is dilute hydrochloric acid, the mass percentage concentration of the solution is 15-25%, when the pickle liquor is dilute sulfuric acid, the mass percentage concentration of the solution is 20-50%, and when the pickle liquor is dilute nitric acid, the mass percentage concentration of the solution is 20-40%.
9. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and S8, monitoring the alkali concentration in the alkali leaching process, and supplementing a certain amount of saturated alkali liquor when the alkali concentration is lower than 15%, so as to ensure that the alkali concentration is more than or equal to 15%.
10. The method for extracting valuable elements from iron ore concentrate by the wet-fire combined process as claimed in claim 1,
and step S8, after alkali treatment, carrying out solid-liquid separation to obtain a titanium-containing solid phase substance and a sodium silicate mother solution, and calcining the titanium-containing solid phase substance at 850-1000 ℃ to obtain a titanium-rich material.
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