CN118145611A - Method for recycling lithium iron phosphate lithium extraction slag - Google Patents
Method for recycling lithium iron phosphate lithium extraction slag Download PDFInfo
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- CN118145611A CN118145611A CN202410306116.1A CN202410306116A CN118145611A CN 118145611 A CN118145611 A CN 118145611A CN 202410306116 A CN202410306116 A CN 202410306116A CN 118145611 A CN118145611 A CN 118145611A
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- slag
- acid
- leaching
- lithium
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- 239000002893 slag Substances 0.000 title claims abstract description 105
- 238000000605 extraction Methods 0.000 title claims abstract description 56
- 238000000034 method Methods 0.000 title claims abstract description 56
- NCZYUKGXRHBAHE-UHFFFAOYSA-K [Li+].P(=O)([O-])([O-])[O-].[Fe+2].[Li+] Chemical compound [Li+].P(=O)([O-])([O-])[O-].[Fe+2].[Li+] NCZYUKGXRHBAHE-UHFFFAOYSA-K 0.000 title abstract description 10
- 238000004064 recycling Methods 0.000 title abstract description 10
- 239000000243 solution Substances 0.000 claims abstract description 78
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 claims abstract description 63
- 239000002253 acid Substances 0.000 claims abstract description 63
- 238000002386 leaching Methods 0.000 claims abstract description 63
- WBJZTOZJJYAKHQ-UHFFFAOYSA-K iron(3+) phosphate Chemical compound [Fe+3].[O-]P([O-])([O-])=O WBJZTOZJJYAKHQ-UHFFFAOYSA-K 0.000 claims abstract description 57
- 229910052782 aluminium Inorganic materials 0.000 claims abstract description 49
- WHXSMMKQMYFTQS-UHFFFAOYSA-N Lithium Chemical compound [Li] WHXSMMKQMYFTQS-UHFFFAOYSA-N 0.000 claims abstract description 48
- 229910052744 lithium Inorganic materials 0.000 claims abstract description 48
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 claims abstract description 45
- 239000003929 acidic solution Substances 0.000 claims abstract description 43
- 239000007787 solid Substances 0.000 claims abstract description 29
- 238000000926 separation method Methods 0.000 claims abstract description 26
- 230000008569 process Effects 0.000 claims abstract description 23
- 229910000398 iron phosphate Inorganic materials 0.000 claims abstract description 17
- 239000003638 chemical reducing agent Substances 0.000 claims abstract description 12
- NBIIXXVUZAFLBC-UHFFFAOYSA-N Phosphoric acid Chemical compound OP(O)(O)=O NBIIXXVUZAFLBC-UHFFFAOYSA-N 0.000 claims description 63
- 239000005955 Ferric phosphate Substances 0.000 claims description 43
- 229940032958 ferric phosphate Drugs 0.000 claims description 43
- 229910000399 iron(III) phosphate Inorganic materials 0.000 claims description 43
- 229910000147 aluminium phosphate Inorganic materials 0.000 claims description 31
- 229940116007 ferrous phosphate Drugs 0.000 claims description 25
- 229910000155 iron(II) phosphate Inorganic materials 0.000 claims description 25
- SDEKDNPYZOERBP-UHFFFAOYSA-H iron(ii) phosphate Chemical compound [Fe+2].[Fe+2].[Fe+2].[O-]P([O-])([O-])=O.[O-]P([O-])([O-])=O SDEKDNPYZOERBP-UHFFFAOYSA-H 0.000 claims description 25
- 238000002156 mixing Methods 0.000 claims description 24
- 238000006243 chemical reaction Methods 0.000 claims description 23
- MHAJPDPJQMAIIY-UHFFFAOYSA-N Hydrogen peroxide Chemical compound OO MHAJPDPJQMAIIY-UHFFFAOYSA-N 0.000 claims description 20
- 238000005245 sintering Methods 0.000 claims description 20
- 239000007853 buffer solution Substances 0.000 claims description 15
- 239000007788 liquid Substances 0.000 claims description 15
- 238000005406 washing Methods 0.000 claims description 15
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 14
- -1 hydrogen ions Chemical class 0.000 claims description 14
- 239000007800 oxidant agent Substances 0.000 claims description 14
- 230000032683 aging Effects 0.000 claims description 13
- GELKBWJHTRAYNV-UHFFFAOYSA-K lithium iron phosphate Chemical compound [Li+].[Fe+2].[O-]P([O-])([O-])=O GELKBWJHTRAYNV-UHFFFAOYSA-K 0.000 claims description 13
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 12
- 238000001035 drying Methods 0.000 claims description 12
- 239000012528 membrane Substances 0.000 claims description 12
- 150000007522 mineralic acids Chemical class 0.000 claims description 11
- 230000001590 oxidative effect Effects 0.000 claims description 9
- 229910052739 hydrogen Inorganic materials 0.000 claims description 8
- 239000001257 hydrogen Substances 0.000 claims description 8
- 239000012266 salt solution Substances 0.000 claims description 8
- GRYLNZFGIOXLOG-UHFFFAOYSA-N Nitric acid Chemical compound O[N+]([O-])=O GRYLNZFGIOXLOG-UHFFFAOYSA-N 0.000 claims description 7
- 239000012452 mother liquor Substances 0.000 claims description 7
- 229910017604 nitric acid Inorganic materials 0.000 claims description 7
- 239000002245 particle Substances 0.000 claims description 7
- 238000001556 precipitation Methods 0.000 claims description 7
- 239000002351 wastewater Substances 0.000 claims description 7
- 238000004321 preservation Methods 0.000 claims description 6
- 235000011121 sodium hydroxide Nutrition 0.000 claims description 4
- ATRRKUHOCOJYRX-UHFFFAOYSA-N Ammonium bicarbonate Chemical compound [NH4+].OC([O-])=O ATRRKUHOCOJYRX-UHFFFAOYSA-N 0.000 claims description 3
- 229910000013 Ammonium bicarbonate Inorganic materials 0.000 claims description 3
- VHUUQVKOLVNVRT-UHFFFAOYSA-N Ammonium hydroxide Chemical compound [NH4+].[OH-] VHUUQVKOLVNVRT-UHFFFAOYSA-N 0.000 claims description 3
- 239000005708 Sodium hypochlorite Substances 0.000 claims description 3
- 235000012538 ammonium bicarbonate Nutrition 0.000 claims description 3
- 239000001099 ammonium carbonate Substances 0.000 claims description 3
- 235000011114 ammonium hydroxide Nutrition 0.000 claims description 3
- 238000010438 heat treatment Methods 0.000 claims description 3
- SUKJFIGYRHOWBL-UHFFFAOYSA-N sodium hypochlorite Chemical compound [Na+].Cl[O-] SUKJFIGYRHOWBL-UHFFFAOYSA-N 0.000 claims description 3
- PFUVRDFDKPNGAV-UHFFFAOYSA-N sodium peroxide Chemical compound [Na+].[Na+].[O-][O-] PFUVRDFDKPNGAV-UHFFFAOYSA-N 0.000 claims description 3
- 239000004411 aluminium Substances 0.000 claims description 2
- 159000000013 aluminium salts Chemical class 0.000 claims description 2
- 229910000329 aluminium sulfate Inorganic materials 0.000 claims description 2
- 239000012535 impurity Substances 0.000 abstract description 30
- 229910052742 iron Inorganic materials 0.000 abstract description 10
- DPTATFGPDCLUTF-UHFFFAOYSA-N phosphanylidyneiron Chemical compound [Fe]#P DPTATFGPDCLUTF-UHFFFAOYSA-N 0.000 abstract description 6
- 230000015572 biosynthetic process Effects 0.000 abstract description 3
- 238000003786 synthesis reaction Methods 0.000 abstract description 3
- 230000007246 mechanism Effects 0.000 abstract description 2
- 239000002699 waste material Substances 0.000 description 13
- 239000000843 powder Substances 0.000 description 9
- 229910019142 PO4 Inorganic materials 0.000 description 8
- 230000000052 comparative effect Effects 0.000 description 8
- 238000001914 filtration Methods 0.000 description 8
- 239000002994 raw material Substances 0.000 description 8
- 230000000694 effects Effects 0.000 description 7
- 239000003513 alkali Substances 0.000 description 6
- 150000004645 aluminates Chemical class 0.000 description 6
- 239000000203 mixture Substances 0.000 description 5
- 229910052698 phosphorus Inorganic materials 0.000 description 5
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 description 4
- 239000011574 phosphorus Substances 0.000 description 4
- 238000007873 sieving Methods 0.000 description 4
- 238000003756 stirring Methods 0.000 description 4
- AZDRQVAHHNSJOQ-UHFFFAOYSA-N alumane Chemical class [AlH3] AZDRQVAHHNSJOQ-UHFFFAOYSA-N 0.000 description 3
- JLDSOYXADOWAKB-UHFFFAOYSA-N aluminium nitrate Inorganic materials [Al+3].[O-][N+]([O-])=O.[O-][N+]([O-])=O.[O-][N+]([O-])=O JLDSOYXADOWAKB-UHFFFAOYSA-N 0.000 description 3
- 150000002500 ions Chemical class 0.000 description 3
- 239000011259 mixed solution Substances 0.000 description 3
- 238000000746 purification Methods 0.000 description 3
- 238000011084 recovery Methods 0.000 description 3
- 238000001878 scanning electron micrograph Methods 0.000 description 3
- CWYNVVGOOAEACU-UHFFFAOYSA-N Fe2+ Chemical compound [Fe+2] CWYNVVGOOAEACU-UHFFFAOYSA-N 0.000 description 2
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 2
- 229910001448 ferrous ion Inorganic materials 0.000 description 2
- 229910052500 inorganic mineral Inorganic materials 0.000 description 2
- 239000011707 mineral Substances 0.000 description 2
- 235000010755 mineral Nutrition 0.000 description 2
- 230000004048 modification Effects 0.000 description 2
- 238000012986 modification Methods 0.000 description 2
- 230000002194 synthesizing effect Effects 0.000 description 2
- 239000010936 titanium Substances 0.000 description 2
- 229910052719 titanium Inorganic materials 0.000 description 2
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 description 2
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 1
- VTLYFUHAOXGGBS-UHFFFAOYSA-N Fe3+ Chemical compound [Fe+3] VTLYFUHAOXGGBS-UHFFFAOYSA-N 0.000 description 1
- RTAQQCXQSZGOHL-UHFFFAOYSA-N Titanium Chemical compound [Ti] RTAQQCXQSZGOHL-UHFFFAOYSA-N 0.000 description 1
- 238000002441 X-ray diffraction Methods 0.000 description 1
- 150000007513 acids Chemical class 0.000 description 1
- 239000010405 anode material Substances 0.000 description 1
- 239000007864 aqueous solution Substances 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 239000010406 cathode material Substances 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 239000003795 chemical substances by application Substances 0.000 description 1
- 239000010949 copper Substances 0.000 description 1
- 229910052802 copper Inorganic materials 0.000 description 1
- 238000005260 corrosion Methods 0.000 description 1
- 230000007797 corrosion Effects 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000007865 diluting Methods 0.000 description 1
- 238000004090 dissolution Methods 0.000 description 1
- 229910001447 ferric ion Inorganic materials 0.000 description 1
- 230000006872 improvement Effects 0.000 description 1
- 230000005764 inhibitory process Effects 0.000 description 1
- BMTOKWDUYJKSCN-UHFFFAOYSA-K iron(3+);phosphate;dihydrate Chemical compound O.O.[Fe+3].[O-]P([O-])([O-])=O BMTOKWDUYJKSCN-UHFFFAOYSA-K 0.000 description 1
- XGZVUEUWXADBQD-UHFFFAOYSA-L lithium carbonate Chemical compound [Li+].[Li+].[O-]C([O-])=O XGZVUEUWXADBQD-UHFFFAOYSA-L 0.000 description 1
- 229910052808 lithium carbonate Inorganic materials 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 238000001728 nano-filtration Methods 0.000 description 1
- 239000002667 nucleating agent Substances 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- 239000010452 phosphate Substances 0.000 description 1
- NBIIXXVUZAFLBC-UHFFFAOYSA-K phosphate Chemical compound [O-]P([O-])([O-])=O NBIIXXVUZAFLBC-UHFFFAOYSA-K 0.000 description 1
- 235000021110 pickles Nutrition 0.000 description 1
- 239000011164 primary particle Substances 0.000 description 1
- 239000012629 purifying agent Substances 0.000 description 1
- 230000035484 reaction time Effects 0.000 description 1
- 239000013557 residual solvent Substances 0.000 description 1
- 150000003839 salts Chemical class 0.000 description 1
- 238000002791 soaking Methods 0.000 description 1
- 239000000126 substance Substances 0.000 description 1
- 239000004094 surface-active agent Substances 0.000 description 1
- 238000004065 wastewater treatment Methods 0.000 description 1
Classifications
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01B—NON-METALLIC ELEMENTS; COMPOUNDS THEREOF; METALLOIDS OR COMPOUNDS THEREOF NOT COVERED BY SUBCLASS C01C
- C01B25/00—Phosphorus; Compounds thereof
- C01B25/16—Oxyacids of phosphorus; Salts thereof
- C01B25/26—Phosphates
- C01B25/37—Phosphates of heavy metals
- C01B25/375—Phosphates of heavy metals of iron
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2002/00—Crystal-structural characteristics
- C01P2002/70—Crystal-structural characteristics defined by measured X-ray, neutron or electron diffraction data
- C01P2002/72—Crystal-structural characteristics defined by measured X-ray, neutron or electron diffraction data by d-values or two theta-values, e.g. as X-ray diagram
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2004/00—Particle morphology
- C01P2004/01—Particle morphology depicted by an image
- C01P2004/03—Particle morphology depicted by an image obtained by SEM
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2004/00—Particle morphology
- C01P2004/60—Particles characterised by their size
- C01P2004/62—Submicrometer sized, i.e. from 0.1-1 micrometer
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2006/00—Physical properties of inorganic compounds
- C01P2006/11—Powder tap density
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2006/00—Physical properties of inorganic compounds
- C01P2006/12—Surface area
-
- C—CHEMISTRY; METALLURGY
- C01—INORGANIC CHEMISTRY
- C01P—INDEXING SCHEME RELATING TO STRUCTURAL AND PHYSICAL ASPECTS OF SOLID INORGANIC COMPOUNDS
- C01P2006/00—Physical properties of inorganic compounds
- C01P2006/80—Compositional purity
Landscapes
- Chemical & Material Sciences (AREA)
- Organic Chemistry (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for recycling lithium iron phosphate lithium extraction slag, and relates to the technical field of lithium batteries. The method comprises the steps of firstly leaching the lithium extraction slag by using a low-concentration first acidic solution to obtain purified ferrophosphorus slag, and controlling the conditions of concentration, liquid-solid ratio and the like of the first acidic solution by using the acid solubility difference of FePO 4 and aluminum, so that the iron in the lithium extraction slag can be better reserved while the aluminum is fully removed; the purified ferrophosphorus slag is mixed and leached with a reducing agent and a high-concentration second acidic solution, fe 3+ in the solution is reduced and H + in the solution is consumed in the leaching process, the pH of the solution is improved, and further impurity removal is realized. The invention utilizes the low-concentration first acid solution and the high-concentration second acid solution to form a double impurity removing mechanism, realizes the separation of impurity aluminum and iron phosphorus in the lithium extraction slag, reduces the impurity content in the subsequent iron phosphate finished product, and can realize the synthesis of the high-purity lamellar iron phosphate product.
Description
Technical Field
The invention relates to the technical field of lithium batteries, in particular to a method for recycling lithium iron phosphate lithium extraction slag.
Background
The lithium iron phosphate battery has the characteristics of excellent electrical circulation performance, high safety performance and low cost, and is applied to the fields of new energy automobiles and the like to be rapidly increased. Iron phosphate is used as a main raw material for preparing lithium iron phosphate, and has important influence on the physical and chemical properties of the lithium iron phosphate. The battery grade iron phosphate used to prepare the cathode material is extremely high in impurity requirements and demanding in basic raw materials, and high-purity phosphoric acid or phosphate and ferric salt are usually used for reaction, however, the cost of the high-purity raw materials is very high. Accordingly, there is a need to consider the use of some low cost sources of iron and phosphorus to synthesize high purity iron phosphate.
The waste residues after extracting lithium from the lithium iron phosphate contain a large amount of valuable ferrophosphorus elements, and can be reprocessed to produce ferric phosphate, so that the waste residues are good low-cost raw materials, but the waste residues also contain impurities such as aluminum and the like, so that the quality of recovered products is affected. Therefore, deep aluminum removal of the ferrophosphorus slag is one of important procedures for recycling the lithium iron phosphate anode material. How to effectively remove aluminum from the ferrophosphorus slag after lithium extraction and synthesize high-purity ferric phosphate is an important difficult problem facing the recovery of the waste lithium iron phosphate battery at present.
The prior ferrophosphorus slag aluminum removal process has the following general problems: (1) The aluminum removal rate is low, and the purity of the subsequent ferric phosphate product is affected; and (2) the process cost is high, and the industrial application is difficult.
In view of this, the present invention has been made.
Disclosure of Invention
The invention aims to provide a method for recycling lithium iron phosphate lithium extraction slag, and aims to provide a simple and easy process for better removing impurity aluminum in the lithium extraction slag.
The invention is realized in the following way:
in a first aspect, the present invention provides a method for recovering lithium iron phosphate lithium extraction slag, comprising:
Mixing and leaching lithium extraction slag obtained after lithium iron phosphate is extracted with a first acidic solution, and then carrying out solid-liquid separation to obtain purified ferrophosphorus slag and aluminum-containing acid liquor; wherein the concentration of hydrogen ions in the first acid solution is 0.1mol/L-0.6mol/L, and the mass ratio of the first acid solution to the lithium extraction slag is (5-15): 1, a step of;
Mixing and leaching the purified ferrophosphorus slag with a second acidic solution and a reducing agent to obtain a ferrous phosphate solution; the concentration of hydrogen ions in the second acidic solution is 1mol/L to 10mol/L.
In an alternative embodiment, when the lithium extraction slag is leached by mixing with the first acidic solution, the leaching temperature is controlled to be 60-90 ℃ and the leaching time is controlled to be 2-10 h;
preferably, the first acidic solution contains an inorganic acid, and the inorganic acid is at least one selected from nitric acid and phosphoric acid;
preferably, the lithium extraction slag is firstly dried and crushed and then mixed and leached with the first acidic solution, and the particle size of the crushed slag is controlled to be 50-100 meshes.
In an alternative embodiment, the aluminium-containing acid solution is subjected to membrane separation to obtain an acid solution to be recovered and an aluminium salt solution;
preferably, the acid solution to be recovered is diluted and then recycled as the first acidic solution.
In an alternative embodiment, the mass ratio of the second acidic solution to the purified ferrophosphorus slag is (12-18): 1, leaching at 35-50 ℃ for 4-8 hours;
preferably, the second acidic solution contains an inorganic acid selected from at least one of sulfuric acid and phosphoric acid.
In an alternative embodiment, the reducing agent is iron powder;
preferably, the molar ratio of the amount of iron powder to Fe 3+ in the purified ferrophosphorus slag is (0.6-1.5): 1.
In an alternative embodiment, in the process of mixing and leaching the purified ferrophosphorus slag with the second acidic solution and the reducing agent, adding a buffer solution to regulate the pH value of the system to be 0.8-2.5;
Preferably, the buffer solution is at least one selected from sodium hydroxide, ammonium bicarbonate and ammonia water, and the mass fraction of the buffer solution is 2% -10%.
In an alternative embodiment, the method further comprises: preparing ferric phosphate by reacting ferrous phosphate solution with oxidant;
Preferably, mixing ferrous phosphate solution with phosphoric acid and oxidant for reaction, controlling the mixing temperature to be 50-70 ℃, mixing for reaction for 20-40min, then heating to 90-110 ℃, preserving heat and aging, and obtaining ferric phosphate solid and precipitation mother liquor after solid-liquid separation;
Preferably, the heat preservation aging time is 6-12 hours;
preferably, the ferrous phosphate solution is heated to 50-70 ℃ and then mixed with phosphoric acid and an oxidant for reaction.
In an alternative embodiment, the molar ratio of Fe 2+ to PO 4 3- in the reaction system is (0.70-0.98): 1 by adjusting the amount of phosphoric acid added.
In an alternative embodiment, the oxidizing agent is selected from at least one of hydrogen peroxide, sodium hypochlorite, and sodium peroxide.
In an alternative embodiment, the method further comprises: washing, drying and sintering the ferric phosphate solid;
Preferably, the iron phosphate solid after washing satisfies: the conductivity of the washing wastewater is less than 500us/cm;
preferably, in the sintering process, the sintering temperature is controlled to be 600-800 ℃ and the sintering time is controlled to be 2-6 h.
The invention has the following beneficial effects: the method comprises the steps of firstly leaching the lithium extraction slag by using a low-concentration first acidic solution to obtain purified ferrophosphorus slag, and controlling the conditions of concentration, liquid-solid ratio and the like of the first acidic solution by using the acid solubility difference of FePO 4 and aluminum, so that the iron in the lithium extraction slag can be better reserved while the aluminum is fully removed; the purified ferrophosphorus slag is mixed and leached with a reducing agent and a high-concentration second acidic solution, fe 3+ in the solution is reduced and H + in the solution is consumed in the leaching process, the pH of the solution is improved, and further impurity removal is realized. The invention utilizes the low-concentration first acid solution and the high-concentration second acid solution to form a double impurity removing mechanism, thereby realizing the separation of impurity aluminum and iron phosphorus in the lithium extraction slag and reducing the impurity content in the subsequent iron phosphate finished product.
Drawings
In order to more clearly illustrate the technical solutions of the embodiments of the present invention, the drawings that are needed in the embodiments will be briefly described below, it being understood that the following drawings only illustrate some embodiments of the present invention and therefore should not be considered as limiting the scope, and other related drawings may be obtained according to these drawings without inventive effort for a person skilled in the art.
FIG. 1 is a flow chart of the purification steps of the method provided in an embodiment of the present invention;
FIG. 2 is an overall process flow diagram for preparing a high purity iron phosphate product from lithium extraction slag;
FIG. 3 is an SEM image of the lithium extraction residue waste material provided in example 1 of the present invention, the synthesized ferric phosphate dihydrate has a magnification of 50000;
Fig. 4 is an SEM image of a 5000-fold magnification of iron phosphate dihydrate synthesized from lithium extraction slag waste provided in example 1 of the present invention;
Fig. 5 is an XRD pattern of the synthesis of anhydrous iron phosphate from lithium extraction residue waste provided in example 1 of the present invention.
Detailed Description
In order to make the objects, technical solutions and advantages of the embodiments of the present invention more clear, the technical solutions of the embodiments of the present invention will be clearly and completely described below. The specific conditions are not noted in the examples and are carried out according to conventional conditions or conditions recommended by the manufacturer. The reagents or apparatus used were conventional products commercially available without the manufacturer's attention.
An embodiment of the invention provides a method for recovering lithium iron phosphate extraction residues, referring to fig. 1 and 2, fig. 1 is a flowchart of purification steps of the method provided by the embodiment of the invention, and fig. 2 is an overall process flow chart for preparing a high-purity iron phosphate product from the extraction residues. The method comprises the following steps:
s1, pretreatment
And taking lithium extraction slag obtained after extracting lithium from lithium iron phosphate, drying the lithium extraction slag, and removing residual solvent on the surface. And then crushing the dried slag, and controlling the particle size of the crushed slag to be 50-100 meshes, so that the slag can be fully contacted with acid liquor in the subsequent acid leaching.
Specifically, the crushed materials can be sieved by a sieve with 50-100 meshes, so that no oversized particles exist. The particle size of the crushed slag can be 50 meshes, 60 meshes, 70 meshes, 80 meshes, 90 meshes, 100 meshes and the like.
The pretreatment step may not be performed, and if the lithium extraction slag raw material itself is dried and the particle size satisfies the requirement, the step may not be performed.
S2, low-concentration acid leaching
And mixing and leaching the lithium extraction slag with a low-concentration first acidic solution, and then carrying out solid-liquid separation to obtain purified ferrophosphorus slag and an aluminum-containing acid solution. By utilizing the acid solubility difference of FePO 4 and aluminum, the conditions of concentration, liquid-solid ratio, leaching temperature, leaching time and the like of the first acid solution are controlled, so that the aluminum removal rate can reach 96%, the iron removal rate is ensured to be lower than 5%, and meanwhile, the iron removal device also has certain impurity removal effect on other impurities (such as Ti, cu and the like), realizes the separation of impurity aluminum and iron phosphorus in waste residues, and reduces the impurity content in the follow-up iron phosphate finished product.
Further, the concentration of hydrogen ions in the first acid solution is 0.1mol/L to 0.6mol/L, and the mass ratio of the first acid solution to the lithium extraction slag is (5-15): the concentration and the amount of the first acidic solution are preferably controlled within the above-mentioned ranges to remove aluminum more sufficiently while retaining iron. If the concentration of the first acidic solution is too high, iron and phosphorus are excessively leached, so that the raw material cannot be efficiently recovered. Specifically, the concentration of hydrogen ions in the first acidic solution may be 0.1mol/L, 0.2mol/L, 0.3mol/L, 0.4mol/L, 0.5mol/L, 0.6mol/L, etc.; the mass ratio (i.e., liquid to solid ratio) of the first acidic solution to the lithium extraction slag may be 5:1, 8:1, 10:1, 12:1, 15:1, etc.
In some embodiments, when the lithium extraction slag is leached by mixing with the first acidic solution, the leaching temperature is controlled to be 60-90 ℃, the leaching time is controlled to be 2-10 h, and the leaching temperature and the leaching time are controlled to be within the above ranges, so that aluminum impurities are more fully removed. Specifically, the leaching temperature at the time of leaching with the first acidic solution may be 60 ℃, 70 ℃, 80 ℃, 90 ℃ and the like, and the leaching time may be 2 hours, 5 hours, 8 hours, 10 hours and the like.
Further, the first acidic solution may be an inorganic acid solution, and the inorganic acid may be at least one selected from nitric acid and phosphoric acid, or may be any one of the above or a mixed solution of two inorganic acids.
In some embodiments, the aluminum-containing acid solution is subjected to membrane separation to obtain an acid solution to be recovered and an aluminum salt solution, and the acid solution to be recovered can be diluted and then recycled as the first acidic solution. The acid solution after selective aluminum extraction is subjected to special separation membrane acid recovery operation, acid liquor to be recovered and aluminum salt solution are filtered and separated, the acid liquor to be recovered is diluted and then is recycled in the step of aluminum removal, and the aluminum salt solution can be used for preparing aluminum water purifying agent, so that the recycling of resources is realized.
Specifically, the separation membrane adopted in membrane separation is not limited, and can be a nanofiltration membrane for short NF, and particularly can be a commercially available NF membrane resistant to acid and alkali, so that the NF membrane can not only resist acid and alkali soaking for a long time, but also resist high-pressure environment, and the running stability is ensured. Diluting the separated acid liquor to be recovered to the concentration of H + of 0.1mol/L-0.6mol/L, and then using the acid liquor to be recovered for low-concentration acid leaching and continuing to use the acid liquor for aluminum extraction.
S3, high-concentration acid leaching
And mixing and leaching the purified ferrophosphorus slag with a reducing agent and a high-concentration second acidic solution to obtain a ferrous phosphate solution. In the leaching process, fe 3+ in the solution is reduced, H + in the solution is consumed, the pH of the solution is improved, further impurity removal is realized, and the removal rate of aluminum is further improved.
Further, the concentration of hydrogen ions in the second acidic solution is 1mol/L to 10mol/L, and the mass ratio of the second acidic solution to the purified ferrophosphorus slag is (12-18): 1, the leaching temperature is 35-50 ℃ and the leaching time is 4-8 h. The concentration, the liquid-solid ratio, the leaching temperature and the leaching time of the second acid solution are further controlled, so that the impurity removal effect is further improved, and the purity of the product is improved.
Specifically, the concentration of hydrogen ions in the second acidic solution may be 1mol/L, 3mol/L, 5mol/L, 8mol/L, 10mol/L, or the like; the leaching temperature can be at normal temperature, such as 15 ℃, 20 ℃, 25 ℃, 30 ℃ and the like, and the leaching time can be 4 hours, 5 hours, 6 hours, 7 hours, 8 hours and the like.
In some embodiments, the second acidic solution may be a mineral acid solution, and the mineral acid is at least one selected from sulfuric acid and phosphoric acid, and may be any one of the above, or may be a mixed solution of two acids.
In some embodiments, the reducing agent is iron powder and the iron powder is used as the reducing agent to avoid the introduction of other impurities and to enable the Fe 3+ to be fully reduced. The molar ratio of the amount of the iron powder to Fe 3+ in the purified ferrophosphorus slag is (0.6-1.5): 1 to fully reduce Fe 3+ to Fe 2+. Specifically, the molar ratio of the amount of iron powder to the Fe 3+ in the purified ferrophosphorus slag may be 0.6:1, 0.8:1, 1.0:1, 1.2:1, 1.5:1, etc.
In some embodiments, during the process of mixing and leaching the purified ferrophosphorus slag with the second acidic solution and the reducing agent, a buffer solution is added to regulate the pH value of the system to be 0.8-2.5 (for example, the pH value can be 0.8, 1.0, 1.2, 1.5, 1.8, 2.0, 2.2, 2.5, etc.), the pH value of the solution is increased by introducing the buffer solution, so that impurities are more fully removed, and the addition amount of the buffer solution in the subsequent synthesis process is reduced.
Further, the buffer solution is at least one selected from sodium hydroxide, ammonium bicarbonate and ammonia water, and can be any one or more of the above, and the mass fraction of the buffer solution is 2% -10%, for example, can be 2%, 5%, 8%, 10% and the like.
S4, preparing ferric phosphate
The ferrous phosphate solution is reacted with an oxidant to prepare ferric phosphate, and the impurity content in the ferrous phosphate solution is reduced by the purification steps of S2 and S3, so that the purity of the prepared ferric phosphate product is improved.
In some embodiments, mixing ferrous phosphate solution with phosphoric acid and oxidant for reaction, controlling the mixing temperature to be 50-70 ℃, mixing for reaction for 20-40min, then heating to 90-110 ℃, preserving heat and aging, and obtaining ferric phosphate solid and precipitation mother liquor after solid-liquid separation. In the reaction process, ferrous ions are oxidized to generate ferric ions, ferric phosphate solid is fully deposited by heat preservation and aging at a higher temperature, and the ferric phosphate solid is obtained after solid-liquid separation for standby.
Specifically, the reaction temperature of the ferrous phosphate solution with phosphoric acid and the oxidizing agent may be 50 ℃, 55 ℃, 60 ℃, 65 ℃, 70 ℃ and the like, and the mixing reaction time may be 20min, 30min, 40min and the like. After the reaction is completed, the temperature is increased for heat preservation and aging, and the temperature for heat preservation and aging can be 90 ℃, 95 ℃, 100 ℃, 105 ℃, 110 ℃ and the like. In order to enable the ferric phosphate solid to be fully deposited, the heat preservation aging time is 6-12 h, such as 6h, 8h, 10h, 12h and the like. Because the pH of the reaction system is low, the reaction is incomplete when the mixture is carried out at low temperature, the driving force is insufficient, and the ferrous phosphate solution can be heated to 50-70 ℃ and then mixed with phosphoric acid and oxidant for reaction.
In some embodiments, the molar ratio of Fe 2+ to PO 4 3- in the reaction system is (0.70-0.98): 1 by adjusting the addition amount of phosphoric acid, i.e., phosphoric acid is added according to the molar ratio of Fe 2+ to PO 4 3- being (0.70-0.98): 1. Specifically, after phosphoric acid is added, the molar ratio of Fe 2+ to PO 4 3- in the system may be 0.70:1, 0.75:1, 0.80:1, 0.85:1, 0.90:1, 0.95:1, 0.98:1, etc.
In some embodiments, the oxidizing agent is at least one selected from hydrogen peroxide, sodium hypochlorite and sodium peroxide, and may be any one or more of the above, and the amount of the oxidizing agent is sufficient to enable sufficient oxidation of ferrous ions.
In the process of synthesizing the ferric phosphate, the cost of raw materials is low, a large amount of alkali is not needed for adjusting the pH value, a nucleating agent is not needed to be added, the production cost of the process is reduced, and the subsequent wastewater treatment procedures are reduced.
S5, post-treatment
And washing, drying and sintering the ferric phosphate solid, and obtaining a sheet ferric phosphate product after post-treatment. The iron phosphate synthesized by the embodiment of the invention has a lamellar structure, the specific surface area after sintering is 4-10m 2/g, the primary particle size is 150-300nm, the tap density is 1.02-1.15g/cm 3, and the purity can reach 99.9%.
The method for removing the impurity ions and unreacted ions remained on the surface of the ferric phosphate solid by washing is not limited, and can be water washing; the washing times are not limited, and the electric conductivity of the washing wastewater is less than 500us/cm, so that the impurity ions on the surface can be removed sufficiently.
In some embodiments, the sintering temperature is controlled to be 600-800 ℃ and the sintering time is controlled to be 2-6 h in the sintering process, and the anhydrous ferric phosphate product can be obtained after sintering. Specifically, the sintering temperature may be 600 ℃, 650 ℃, 700 ℃, 750 ℃, 800 ℃, etc.; the sintering time can be 2h, 3h, 4h, 5h, 6h, etc.
The features and capabilities of the present invention are described in further detail below in connection with the examples.
It should be noted that, the lithium extraction slag used in the following examples and comparative examples is obtained after lithium iron phosphate is extracted, and specific lithium extraction process can refer to CN115744940a, hydrochloric acid and hydrogen peroxide are used as leaching agents, lithium is selectively leached, and surfactant is added to strengthen the leaching process of waste lithium iron phosphate positive electrode powder; removing impurities from the leaching solution, concentrating, and synthesizing to prepare lithium carbonate.
Example 1
The embodiment provides a method for recycling lithium iron phosphate lithium extraction slag, which comprises the following steps:
(1) Drying and crushing lithium extraction slag, and sieving the dried and crushed lithium extraction slag with a 50-mesh sieve to obtain slag powder.
(2) Mixing the waste residue powder with a mixed acid solution formed by HNO 3 and H 3PO4, wherein the total concentration of H + in the mixed acid solution is 0.3mol/L, and the molar ratio of HNO 3 to H + in H 3PO4 is 3:1, controlling the liquid-solid ratio to be 8:1, the reaction temperature is 60 ℃, the leaching time is 3 hours, then solid-liquid separation is carried out to obtain low aluminum phosphorus iron slag and aluminate containing the aluminum, and acid-alkali-resistant NF membrane separation is utilized to recycle nitric acid and aluminum salt solution in the aluminate containing the aluminum.
(3) Adding the low-aluminum iron phosphate slag into a mixed acid solution, and adding phosphoric acid into the mixed acid: sulfuric acid = 2:1 (molar ratio, same below), total H + concentration is 4mol/L, meanwhile, iron powder is added, the molar ratio of the iron powder to Fe 3+ in the ferrophosphorus slag is 0.6:1, stirring and filtering are carried out, buffer solution (sodium hydroxide aqueous solution with mass fraction of 5% and the same below) is added, pH value of the solution is controlled to be 0.8-1.5, leaching is carried out for 5 hours under the condition of 45 ℃ to obtain ferrous phosphate solution.
(4) And (3) raising the temperature of the ferrous phosphate solution to 50 ℃, adding phosphoric acid into the ferrous phosphate solution to adjust the Fe/P molar ratio to 0.95, adding hydrogen peroxide with the concentration of 30% which is 1.2 times of the molar ratio, reacting for 1h, raising the temperature to 90 ℃, preserving heat and aging for 10h, and carrying out solid-liquid separation on the product to obtain ferric phosphate solid and precipitation mother liquor.
(5) Washing the ferric phosphate solid until the conductivity of the wastewater is less than 500us/cm, filtering, drying to obtain a flaky ferric phosphate dihydrate product, and sintering at 650 ℃ for 6 hours to obtain anhydrous ferric phosphate.
The compositions of the low-aluminum ferrophosphorus slag obtained in the step (2) and the leachate and leaching slag obtained in the step (3) of this example were tested, and the results are shown in table 1.
TABLE 1 variation of the content of the elements during the process
Fe | Al | P | Cu | Ti | |
Lithium extraction slag | 27.36% | 0.75% | 16.82% | 0.039% | 0.31% |
Low aluminium phosphorus iron slag | 26.92% | 0.061% | 16.31% | 0.032% | 0.28% |
Leachate solution | 92.5g/L | 12mg/L | 62.7g/L | 15mg/L | 11mg/L |
Leaching residue | 0.71% | 0.041% | 0.52% | 0.03% | 0.25% |
Example 1 SEM images of the synthesized ferric phosphate dihydrate are shown in fig. 3 and 4, and it can be seen that the prepared ferric phosphate dihydrate particles are uniform and lamellar. Example 1 XRD test patterns of the synthesized ferric phosphate dihydrate are shown in fig. 5, from which conclusions can be drawn about the synthesized ferric phosphate dihydrate by comparison with standard cards.
Example 2
The embodiment provides a method for recycling lithium iron phosphate lithium extraction slag, which comprises the following steps:
(1) Drying and crushing lithium extraction slag, and sieving the dried and crushed lithium extraction slag with a 50-mesh sieve to obtain slag powder.
(2) Mixing the waste residue powder with 0.3mol/L HNO 3 solution, and controlling the liquid-solid ratio to be 10:1, the reaction temperature is 70 ℃ and the leaching time is 4 hours, then the low-aluminum phosphorus iron slag and the aluminate containing the aluminum are obtained after solid-liquid separation, and nitric acid and aluminum salt solution in the aluminate containing the aluminum are recovered by utilizing acid-alkali resistant NF membrane separation.
(3) Adding the low-aluminum iron phosphate slag into a mixed acid solution, and adding phosphoric acid into the mixed acid: sulfuric acid = 2:1, the total H + concentration is 4mol/L, meanwhile, iron powder is added, the mole ratio of the iron powder to Fe 3+ in the ferrophosphorus slag is 0.6:1, stirring and filtering are carried out, buffer solution is added, the pH value of the solution is controlled to be 0.8-1.5, and leaching is carried out for 5 hours under the condition of 45 ℃ to obtain ferrous phosphate solution.
(4) And (3) raising the temperature of the ferrous phosphate solution to 50 ℃, adding phosphoric acid to adjust the molar ratio of Fe/P to 0.95, adding hydrogen peroxide with the concentration of 30% which is 1.2 times of the molar ratio of Fe/P, reacting for 1h, raising the temperature to 90 ℃, preserving heat and aging for 10h, and carrying out solid-liquid separation on the product to obtain ferric phosphate solid and precipitation mother liquor.
(5) Washing the ferric phosphate solid until the conductivity of the wastewater is less than 500us/cm, filtering, drying to obtain a flaky ferric phosphate dihydrate product, and sintering at 650 ℃ for 6 hours to obtain anhydrous ferric phosphate.
The compositions of the low-aluminum ferrophosphorus slag obtained in the step (2) and the leachate and leaching slag obtained in the step (3) of this example were tested, and the results are shown in Table 2.
Table 2: the content of each element changes in the process
Fe | Al | P | Cu | Ti | |
Lithium extraction slag | 27.36% | 0.75% | 16.82% | 0.039% | 0.31% |
Low aluminium phosphorus iron slag | 26.15% | 0.030% | 15.75% | 0.029% | 0.28% |
Leachate solution | 92.4g/L | 19mg/L | 63.4g/L | 18mg/L | 12mg/L |
Leaching residue | 0.74% | 0.026% | 0.40% | 0.025% | 0.20% |
Example 3
The embodiment provides a method for recycling lithium iron phosphate lithium extraction slag, which comprises the following steps:
(1) Drying and crushing lithium extraction slag, and sieving the dried and crushed lithium extraction slag with a 50-mesh sieve to obtain slag powder.
(2) Mixing the waste residue powder with 0.15mol/L H 3PO4 solution, and controlling the liquid-solid ratio to be 15:1, the reaction temperature is 80 ℃, the leaching time is 10 hours, the low aluminum phosphorus iron slag and the aluminate containing the aluminum are obtained after solid-liquid separation, and nitric acid and aluminum salt solution in the aluminate containing the aluminum are recovered by utilizing acid-alkali resistant NF membrane separation.
(3) Adding the low-aluminum iron phosphate slag into a mixed acid solution, and adding phosphoric acid into the mixed acid: sulfuric acid = 2:1, the total H + concentration is 4mol/L, and meanwhile, iron powder is added, and the molar ratio of the iron powder to Fe 3+ in the ferrophosphorus slag is 0.6:1, stirring and filtering, adding a buffer solution, controlling the pH of the solution to be 0.8-1.5, and leaching for 5 hours at 45 ℃ to obtain a ferrous phosphate solution.
(4) And (3) raising the temperature of the ferrous phosphate solution to 50 ℃, adding phosphoric acid to adjust the Fe/P molar ratio to 0.95, adding hydrogen peroxide with the concentration of 30% which is 1.2 times of the molar ratio, reacting for 1h, raising the temperature to 90 ℃, preserving heat and aging for 10h, and carrying out solid-liquid separation on the product to obtain ferric phosphate solid and precipitation mother liquor.
(5) Washing the ferric phosphate solid until the conductivity of the wastewater is less than 500us/cm, filtering, drying to obtain a flaky ferric phosphate dihydrate product, and sintering at 650 ℃ for 6 hours to obtain anhydrous ferric phosphate.
The compositions of the low-aluminum ferrophosphorus slag obtained in the step (2) and the leachate and leaching slag obtained in the step (3) of this example were tested, and the results are shown in Table 3.
Table 3: the content of each element changes in the process
Fe | Al | P | Cu | Ti | |
Lithium extraction slag | 27.36% | 0.75% | 16.82% | 0.039% | 0.31% |
Low aluminium phosphorus iron slag | 26.42% | 0.19% | 16.42% | 0.03% | 0.22% |
Leachate solution | 90.6g/L | 26mg/L | 63.3g/L | 12mg/L | 7mg/L |
Leaching residue | 0.68% | 0.025% | 0.45% | 0.026% | 0.20% |
It should be noted that example 2 and example 3 differ from example 1 only in that in step (2), the effect of different acid raw materials and corresponding leaching conditions on the impurity removal effect can be compared.
Example 4
The only difference from example 1 is that: the ratio of the iron powder in the step (3) to the Fe 3+ in the ferrophosphorus slag is 1.
Example 5
The only difference from example 1 is that: the ratio of the iron powder in the step (3) to the Fe 3+ in the ferrophosphorus slag is 1.5.
Example 6
The only difference from example 1 is that: and (3) adding phosphoric acid in the step (4) to adjust the iron-phosphorus ratio to 0.9.
Example 7
The only difference from example 1 is that: and (3) adding phosphoric acid in the step (4) to adjust the iron-phosphorus ratio to 0.8.
Example 8
The only difference from example 1 is that: in the step (4), the temperature of the ferrous phosphate solution is increased to 60 ℃, and then equal amount of phosphoric acid and hydrogen peroxide are added for reaction.
Example 9
The only difference from example 1 is that: and (4) raising the temperature of the ferrous phosphate solution to 70 ℃, and then adding the same amount of phosphoric acid and hydrogen peroxide for reaction.
Examples 10 to 13
The only difference from example 1 is that: the concentration of H + in the total acid used in the step (2) was varied, and the concentration of H + in examples 10 to 13 was 0.1mol/L, 0.4mol/L, 0.5mol/L, and 0.6mol/L in this order.
Comparative example 1
This comparative example provides a method for recovering lithium iron phosphate lithium extraction slag, providing a conventional recovery method, comprising the steps of:
(1) Drying and crushing lithium extraction slag, and sieving the dried and crushed lithium extraction slag with a 50-mesh sieve to obtain slag powder.
(2) Directly adding the waste residue powder into a mixed acid solution, wherein phosphoric acid in the mixed acid is as follows: sulfuric acid = 2:1, the total H + concentration is 4mol/L, meanwhile, iron powder is added, the ratio of the iron powder to Fe 3+ in the ferrophosphorus slag is 0.6, stirring and filtering are carried out, buffer solution is added, the pH value of the solution is controlled to be 0.8-1.5, and leaching is carried out for 5 hours under the condition of 45 ℃ to obtain ferrous phosphate solution.
(3) And (3) raising the temperature of the ferrous phosphate solution to 50 ℃, adding phosphoric acid to adjust the molar ratio of Fe/P to 0.95, adding hydrogen peroxide with the concentration of 30% which is 1.2 times of the molar ratio of Fe/P, reacting for 1h, raising the temperature to 90 ℃, preserving heat and aging for 10h, and carrying out solid-liquid separation on the product to obtain ferric phosphate solid and precipitation mother liquor.
(4) Washing the ferric phosphate solid until the conductivity of the wastewater is less than 500us/cm, filtering, drying to obtain a flaky ferric phosphate dihydrate product, and sintering at 650 ℃ for 6 hours to obtain anhydrous ferric phosphate.
The composition of the leachate and the leaching residue obtained in the step (2) of this comparative example was tested, and the results are shown in Table 4.
Table 4: the content of each element changes in the process
Fe | Al | P | Cu | Ti | |
Lithium extraction slag | 27.36% | 0.75% | 16.82% | 0.039% | 0.31% |
Leachate solution | 90.3g/L | 646mg/L | 61.6g/L | 51mg/L | 29mg/L |
Leaching residue | 0.72% | 0.32% | 0.74% | 0.021% | 0.27% |
Comparative examples 2 to 3
The only difference from example 1 is that: the concentration of H + in the total acid used in the step (2) was varied, and the concentration of H + in the comparative examples 2 to 3 was 2mol/L and 5.0mol/L in this order.
Test example 1
The parameters of the iron phosphate product of the above example were tested according to the HG/T4701-2021 standard, and the relevant product properties are shown in Table 5.
Table 5: content of each impurity element, fe/P, BET and tap density of synthesized ferric phosphate product
Comparative examples 1 to 3 did not undergo the step of removing aluminum by using a low concentration mixed acid, but the iron powder was directly added with a high concentration mixed acid to dissolve out iron and phosphorus, and it can be seen from table 5 that the impurity contents of aluminum, copper and titanium in the finally obtained leachate were all higher, the purity of the finally synthesized product was also lower, and the higher the acid concentration was, the higher the content of dissolved out impurity was, and the more impure the synthesized product was. In addition to comparative examples 1-3, the examples of the steps of adding low-concentration mixed acid and removing aluminum are added, the purity of the final synthesized product is higher, and the Al content in the product is lower than 0.0010 percent and meets the standard.
In examples 10 to 13, the total H + concentration of the low-concentration acid leaching is changed, when the H + concentration is high, the loss rate of iron and phosphorus in the ferrophosphorus slag is also high, and when the concentration is low, the removal rate of aluminum is low, but a second aluminum removal process is also present later, so that the influence on the product is small.
In comparison with examples 1-3, at the same H + concentration, the primary aluminum removal effect HNO 3>HNO3 and H 3PO4 are mixed with acid > H 3PO4, because FePO 4 is slightly soluble in inorganic acid, the solubility of HNO 3>H3PO4 in acid, and Fe 3+ in pickle liquor forms a corrosion battery with Al, thereby improving the dissolution rate of aluminum. HNO 3 has the best effect on removing aluminum impurities in the lithium extraction slag, but meanwhile, the loss rate of Fe and P in the slag is higher; the use of H 3PO4 has an inhibition effect on the leaching of FePO 4, but is difficult to leach impurity aluminum, and the liquid-solid ratio is required to be larger, the temperature is higher, and the leaching time is longer, so that the cost is increased; the mixed solution of HNO 3 and H 3PO4 can combine the advantages of the two acid leaching methods, has optimal selectivity to aluminum and ferrophosphorus, can remove more aluminum on the basis of least loss of ferrophosphorus, and has more proper leaching time and lower cost.
The above is only a preferred embodiment of the present invention, and is not intended to limit the present invention, but various modifications and variations can be made to the present invention by those skilled in the art. Any modification, equivalent replacement, improvement, etc. made within the spirit and principle of the present invention should be included in the protection scope of the present invention.
Claims (10)
1. A method for recovering lithium iron phosphate extracted lithium slag, which is characterized by comprising the following steps:
Mixing and leaching lithium extraction slag obtained after lithium iron phosphate is extracted with a first acidic solution, and then carrying out solid-liquid separation to obtain purified ferrophosphorus slag and aluminum-containing acid liquor; wherein the concentration of hydrogen ions in the first acidic solution is 0.1mol/L-0.6mol/L, and the mass ratio of the first acidic solution to the lithium extraction slag is (5-15): 1, a step of;
Mixing and leaching the purified ferrophosphorus slag with a second acidic solution and a reducing agent to obtain a ferrous phosphate solution; the concentration of hydrogen ions in the second acidic solution is 1mol/L-10mol/L.
2. The method according to claim 1, wherein when the lithium extraction slag is leached by mixing with the first acidic solution, the leaching temperature is controlled to be 60-90 ℃ and the leaching time is controlled to be 2-10 h;
Preferably, the first acidic solution contains an inorganic acid, and the inorganic acid is at least one selected from nitric acid and phosphoric acid;
Preferably, the lithium extraction slag is firstly dried and crushed and then mixed and leached with the first acidic solution, and the particle size of the crushed slag is controlled to be 50-100 meshes.
3. The method according to claim 2, wherein the aluminium-containing acid solution is subjected to membrane separation to obtain an acid solution to be recovered and an aluminium salt solution;
preferably, the acid solution to be recovered is diluted and then recycled as the first acidic solution.
4. The method according to claim 1, wherein the mass ratio of the second acidic solution to the purified ferrophosphorus slag is (12-18): 1, leaching at 35-50 ℃ for 4-8 hours;
Preferably, the second acidic solution contains an inorganic acid, and the inorganic acid is at least one selected from sulfuric acid and phosphoric acid.
5. The method of claim 4, wherein the reducing agent is iron powder;
Preferably, the molar ratio of the amount of the iron powder to the Fe 3+ in the purified ferrophosphorus slag is (0.6-1.5): 1.
6. The method according to claim 4 or 5, wherein during the process of mixed leaching of the purified ferrophosphorus slag with the second acidic solution and the reducing agent, a buffer solution is added to regulate the pH value of the system to be 0.8-2.5;
Preferably, the buffer solution is at least one selected from sodium hydroxide, ammonium bicarbonate and ammonia water, and the mass fraction of the buffer solution is 2% -10%.
7. The method as recited in claim 1, further comprising: preparing ferric phosphate by utilizing the ferrous phosphate solution to react with an oxidant;
Preferably, mixing the ferrous phosphate solution with phosphoric acid and an oxidant for reaction, controlling the mixing temperature to be 50-70 ℃, mixing for reaction for 20-40min, then heating to 90-110 ℃, preserving heat and aging, and obtaining ferric phosphate solid and precipitation mother liquor after solid-liquid separation;
Preferably, the heat preservation aging time is 6-12 hours;
Preferably, the ferrous phosphate solution is heated to 50-70 ℃ and then mixed with phosphoric acid and the oxidant for reaction.
8. The method according to claim 7, wherein the molar ratio of Fe 2+ to PO 4 3- in the reaction system is (0.70-0.98): 1 by adjusting the amount of phosphoric acid added.
9. The method of claim 7, wherein the oxidizing agent is selected from at least one of hydrogen peroxide, sodium hypochlorite, and sodium peroxide.
10. The method as recited in claim 7, further comprising: washing, drying and sintering the ferric phosphate solid;
preferably, the iron phosphate solid after washing satisfies: the conductivity of the washing wastewater is less than 500us/cm;
preferably, in the sintering process, the sintering temperature is controlled to be 600-800 ℃ and the sintering time is controlled to be 2-6 h.
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