Method for recovering metal from crude nickel cobalt hydroxide

Abstract

The invention relates to the technical field of metal recovery, in particular to a method for recovering metal from coarse nickel cobalt hydroxide. The method for recycling the metal from the coarse nickel cobalt hydroxide comprises the following steps that S1, after the coarse nickel cobalt hydroxide is subjected to ammonia complexing leaching and oxidation in sequence, leachate and manganese slag are obtained; s2, after the leachate is separated through an extraction method, raffinate, a nickel-containing solution and a copper-containing solution are obtained; s3, the raffinate is subjected to ammonia distillation treatment, and cobalt-zinc slag is obtained; s4, the cobalt-zinc slag is subjected to acid leaching, and a zinc-containing solution and cobalt slag are obtained; and S5, the cobalt slag is subjected to reduction leaching, and a cobalt-containing solution is obtained. According to the method for recovering the metal from the crude nickel cobalt hydroxide, the metal in the crude nickel cobalt hydroxide can be fully extracted, and a nickel product, a cobalt product, a copper product, a zinc product and manganese slag which can be sold as a manganese smelting raw material are obtained.

Description

Technical Field

[0001] The invention relates to the technical field of metal recovery, in particular to a method for recovering metals from crude nickel and cobalt hydroxide. Background Art

[0002] Crude nickel cobalt hydroxide (MHP) is one of the main raw materials for nickel hydrometallurgy. Compared with other nickel raw materials, it has the advantages of rapid dissolution and low process requirements. At the same time, as an alkaline material, it can be used as a pH regulator. At present, the main process for processing MHP in industrial production is: reduction acid leaching, chemical iron removal and extraction and impurity separation in sequence to prepare nickel sulfate solution, cobalt sulfate solution or corresponding crystals.

[0003] During the reduction acid leaching process, the metal elements will basically react with the acid and enter the leachate. The leachate has many impurities and needs to be treated by chemical iron removal, P204 extraction and impurity removal, P507 extraction to prepare cobalt sulfate, and C272 magnesium removal to finally obtain a nickel sulfate solution. The process is complicated, and the cost of auxiliary materials consumed in solution purification and impurity removal is high. At the same time, extraction wastewater and waste residue are generated, which increases the pressure on production and environmental protection and has high production energy consumption. Chemical iron removal is the chemical precipitation of iron into iron alum, which removes most of the iron and has no utilization value and may become hazardous waste.

[0004] The P204 extraction and impurity removal process removes Mn, Ca, Cu, etc. in the form of the so-called "copper-manganese liquid", and the Mn and Cu are subsequently specially processed and recovered. In this process, Zn is discharged in the iron washing liquid together with a trace amount of iron, and is precipitated into solid slag after sewage treatment, which has no actual recovery value. The preparation of cobalt sulfate by P507 extraction is to extract the cobalt from the P204 extraction residue, separate it from the nickel and magnesium therein, and obtain a cobalt sulfate solution. C272 magnesium removal is to extract the magnesium from the P507 residue, separate it from the nickel therein, and obtain a magnesium sulfate solution; a small amount of nickel will be entrained into the magnesium sulfate solution in this process, resulting in nickel loss. This part of the solution also needs to be treated with sewage and precipitated into solid slag, which has no actual recovery value. The residual liquid from the C272 extraction is nickel sulfate liquid. Although other impurity metal elements are basically separated out, the anions are not removed. If the types of anions brought by the raw materials are too many (such as phosphorus, chlorine, silicon, etc.) or the content is too high, they cannot be directly used in the ternary precursor, but must be crystallized into crystals before use, which invisibly increases the production cost.

[0005] Due to the complex composition of the leachate obtained by reduction acid leaching, the impurity removal and purification process is long and only two qualified products, cobalt sulfate and nickel sulfate, can be obtained. Other valuable metals such as Mn, Cu, Zn, etc. must be prepared by other supporting processes, which is costly and inefficient.

[0006] In view of this, this invention is proposed. Summary of the invention

[0007] The object of the present invention is to provide a method for recovering metals from crude nickel cobalt hydroxide, which can fully extract the metals in the crude nickel cobalt hydroxide to obtain nickel products, cobalt products, copper products, zinc products and manganese slag.

[0008] In order to achieve the above-mentioned purpose of the present invention, the following technical solutions are particularly adopted:

[0009] The present invention provides a method for recovering metals from crude nickel and cobalt hydroxide, comprising the following steps:

[0010] S1, subjecting the crude nickel and cobalt hydroxide to ammonia complex leaching and oxidation in sequence to obtain a leachate and manganese slag;

[0011] S2, the leachate is separated by extraction to obtain a raffinate, a nickel-containing solution and a copper-containing solution;

[0012] S3, subjecting the raffinate to ammonia evaporation treatment to obtain cobalt-zinc slag;

[0013] S4, acid leaching the cobalt-zinc slag to obtain a zinc-containing solution and cobalt slag;

[0014] S5, reducing and leaching the cobalt slag to obtain a cobalt-containing solution.

[0015] Furthermore, in step S1, the ammonia complex leaching includes: adding ammonium salt and ammonia water to the slurry obtained after slurrying of crude nickel cobalt hydroxide, and leaching at below 50° C. for 4 to 6 hours.

[0016] Furthermore, in step S1, the ammonium salt includes ammonium bicarbonate and / or ammonium carbonate.

[0017] Further, in step S1, the mass ratio of the crude nickel cobalt hydroxide to the ammonium salt is 100:(65-80);

[0018] And / or, the usage ratio of the crude nickel cobalt hydroxide and the ammonia water is 100g:400-600mL.

[0019] Furthermore, in step S1, the oxidation includes: introducing air into the slurry after the ammonia complex leaching for oxidation.

[0020] Furthermore, in step S2, the extraction separation comprises: extracting the leachate with Lix84 extractant to obtain an extracted organic phase and a raffinate; and subjecting the extracted organic phase to secondary stripping to obtain the nickel-containing solution and the copper-containing solution.

[0021] Further, in step S2, the extracted organic phase is subjected to a first back extraction with sulfuric acid to obtain the nickel-containing solution and the first organic phase, and the pH of the nickel-containing solution is 1.5-2.0; the first organic phase is subjected to a second back extraction with sulfuric acid to obtain the copper-containing solution, and the acidity of the copper-containing solution is 1.8-2.2N.

[0022] Furthermore, in step S4, the acid leaching comprises: adding a sulfuric acid solution to the cobalt-zinc slag, and performing acid leaching at a pH of 1 to 1.5 to obtain the zinc-containing solution and the cobalt slag.

[0023] Further, in step S5, the reduction leaching comprises: adding sulfuric acid and a reducing agent to the cobalt slag, and performing reduction leaching at a pH of 0.5 to 1.5 to obtain the cobalt-containing solution.

[0024] Furthermore, in step S5, the reducing agent includes at least one of hydrogen peroxide, sodium sulfite and sodium metabisulfite.

[0025] Compared with the prior art, the present invention has the following beneficial effects:

[0026] The method for recovering metals from MHP of the present invention comprises leaching metals Ni, Co, Cu, Zn, Mn from MHP by ammonia complex leaching. 2+ A small amount of Mn will also be leached out, and the Mn in the solution after leaching by oxidative ammonia complexation 2+ and Co 2+ , so that Mn is completely retained in the slag to obtain manganese slag, achieving the purpose of completely separating manganese, Co 2+ Oxidized to Co 3+ ; Then use Lix84 extractant, Zn 2+ and Co 3+ During the extraction, it is not extracted but remains in the raffinate, thus being separated from Ni and Cu. The raffinate is treated with ammonia evaporation. After the ammonia is evaporated, Co 3+ and Zn 2+ It turns into hydroxide and precipitates to form cobalt-zinc slag; using Co 3+ The method for recovering metals in MHP of the present invention can fully extract the metals in the crude nickel and cobalt hydroxide to obtain nickel products, cobalt products, copper products, zinc products and manganese slag; and there is no waste water and waste slag discharged, and zero discharge can be achieved. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] In order to more clearly illustrate the technical solutions of the embodiments of the present invention, the drawings required for use in the embodiments will be briefly introduced below. It should be understood that the following drawings only show certain embodiments of the present invention and therefore should not be regarded as limiting the scope. For ordinary technicians in this field, other related drawings can be obtained based on these drawings without paying creative work.

[0028] Figure 1 The figure is a schematic flow diagram of the method for recovering metals in MHP of the present invention. DETAILED DESCRIPTION

[0029] The technical scheme of the present invention will be clearly and completely described below in conjunction with the accompanying drawings and specific embodiments, but it will be understood by those skilled in the art that the following described embodiments are part of embodiments of the present invention, rather than all embodiments, and are only used to illustrate the present invention, and should not be considered as limiting the scope of the present invention. Based on the embodiments in the present invention, all other embodiments obtained by those of ordinary skill in the art without making creative work are within the scope of protection of the present invention. If specific conditions are not specified in the embodiments, they are carried out according to normal conditions or conditions recommended by the manufacturer. If the manufacturer is not specified in the reagents or instruments used, they are all conventional products that can be purchased commercially.

[0030] The following is a specific description of the method for recovering metals from crude nickel and cobalt hydroxide of the present invention.

[0031] See also Figure 1 In some embodiments of the present invention, a method for recovering metals from crude nickel cobalt hydroxide (MHP) is provided, comprising the following steps:

[0032] S1, subjecting the crude nickel and cobalt hydroxide to ammonia complex leaching and oxidation in sequence to obtain a leachate and manganese slag;

[0033] S2, the leachate is separated by extraction to obtain a raffinate, a nickel-containing solution and a copper-containing solution;

[0034] S3, subjecting the raffinate to ammonia evaporation treatment to obtain cobalt-zinc slag;

[0035] S4, acid leaching the cobalt-zinc slag to obtain a zinc-containing solution and cobalt slag;

[0036] S5. Reduction leaching is performed on the cobalt slag to obtain a cobalt-containing solution.

[0037] The method for recovering metals from MHP of the present invention is to leach metals Ni, Co, Cu and Zn from MHP into a solution by ammonia complex leaching, while Mn and other impurities remain in the solid slag; after leaching, due to Mn 2+ A small amount of Mn will also be leached out, so the solution after oxidation leaching is made to enter the solution in a small amount.2+ The solid slag is then separated by extraction to extract Ni and Cu in the solution into the organic phase, and Co 3+ and Zn 2+ The extracted organic phase is then stripped to obtain a nickel-containing solution and a copper-containing solution; the raffinate is treated with ammonia evaporation, and after the ammonia is evaporated, Co 3+ and Zn 2+ The complex ions are lost and become free ions. In an alkaline environment, they become hydroxides and precipitate to form cobalt-zinc slag. After the cobalt-zinc slag is acid-leached, Zn 2+ leached into the solution, while Co 3+ It is not dissolved and remains in the slag to form cobalt slag; the cobalt slag is subjected to reduction leaching to obtain a cobalt-containing solution.

[0038] The method for recovering metals from crude nickel cobalt hydroxide of the present invention can fully extract the metals from the crude nickel cobalt hydroxide to obtain nickel products, cobalt products, copper products, zinc products, and manganese slag. The manganese slag can be sold as a smelting raw material for manganese; and no waste water and waste slag are discharged, so zero emission can be achieved.

[0039] In some embodiments of the present invention, in step S1, ammonia complex leaching includes: adding ammonium salt and ammonia water to the slurry obtained after slurrying of crude nickel cobalt hydroxide, and leaching at below 50°C for 4 to 6 hours; typically but not limitatively, for example, the temperature of ammonia complex leaching can be 45°C, 40°C, 35°C, 30°C, 25°C, 20°C or a range of any two thereof; the time of ammonia complex leaching can be 4h, 5h, 6h or a range of any two thereof; preferably, the solid content of the slurry is 20% to 30%.

[0040] In some embodiments of the present invention, in step S1, the ammonium salt includes ammonium bicarbonate and / or ammonium carbonate.

[0041] In some embodiments of the present invention, in step S1, the mass ratio of crude nickel cobalt hydroxide and ammonium salt is 100:(65-80); typically but not limiting, for example, the mass ratio of crude nickel cobalt hydroxide and ammonium salt can be 100:65, 100:70, 100:75, 100:80 or a range of any two thereof.

[0042] In some embodiments of the present invention, in step S1, the usage ratio of crude nickel cobalt hydroxide and ammonia water is 100g:400-600mL; typically but not limitatively, for example, the usage ratio of crude nickel cobalt hydroxide and ammonia water can be 100g:400mL, 100g:450mL, 100g:500mL, 100g:550mL, 100g:600mL or a range value composed of any two thereof; preferably, the concentration of ammonia water is 25%.

[0043] In some embodiments of the present invention, in step S1, oxidation includes: introducing air into the slurry after ammonia complex leaching for oxidation.

[0044] In some embodiments of the present invention, in step S1, the oxidation temperature is below 50°C, preferably 35-50°C; the oxidation time is 2-5h, preferably 3-4h.

[0045] In step S1 of the present invention, after MHP is slurried with water, ammonia complex leaching is performed to leach the metals Ni, Co, Cu, and Zn in MHP into the solution, but Mn 2+ A small amount of Mn will also be leached (the Mn in the solution obtained after ammonia complex leaching 2+ The content is 0.1-1.7 g / L); therefore, after leaching, air is introduced to remove the Co 2+ and a small amount of Mn 2+ Oxidized to high valence, Co 3+ It forms a complex with ammonia and remains in the solution, and manganese enters the solid slag in the form of MnO(OH) or MnO(OH)2, thus achieving the purpose of completely separating manganese.

[0046] Since Mn(NH3)4 2+ The reduction potential is higher than that of free Mn 2+ Lower, so the Mn is oxidized in the solution obtained after ammonia complex leaching 2+ It is easier. As long as the oxygen in the air is fully mixed with the solution, it can convert Mn(NH3)4 2+ Oxidation causes a small amount of manganese that enters the solution to precipitate into the manganese slag. Co(NH3)4 2+ It is also easily oxidized to Co(NH3)6 3+ ion.

[0047] The chemical reactions that occur during ammonia complex leaching and oxidation are as follows:

[0048] Ni 2+ +4NH3=Ni(NH3)4 2+ ;

[0049] Co 2+ +4NH3=Co(NH3)4 2+ ;

[0050] Cu 2+ +4NH3=Cu(NH3)4 2+ ;

[0051] Zn 2+ +4NH3=Zn(NH3)4 2+ ;

[0052] Mn 2+ +4NH3=Mn(NH3)4 2+ ;

[0053] 4Co(NH3)4 2+ +O2+8NH3·H2O=4Co(NH3)6 3+ +4OH - +6H2O;

[0054] 2Mn(NH3)4 2+ +O2+2OH - =2MnO(OH)↓+8NH3;

[0055] 2Mn(NH3)4 2+ +O2+4OH - =2MnO(OH)2↓+8NH3.

[0056] In some embodiments of the present invention, in step S2, separation by extraction method comprises: extracting the leachate with Lix84 extractant to obtain an extracted organic phase and a raffinate; and subjecting the extracted organic phase to secondary stripping to obtain a nickel-containing solution and a copper-containing solution.

[0057] In some embodiments of the present invention, in step S2, after the extracted organic phase is subjected to a first back extraction with sulfuric acid, a nickel-containing solution and a first organic phase are obtained, and the pH of the nickel-containing solution is 1.5-2.0; after the first organic phase is subjected to a second back extraction with sulfuric acid, a copper-containing solution is obtained, and the acidity of the copper-containing solution is 1.8-2.2N; the nickel-containing solution is a nickel sulfate solution, and the copper-containing solution is a copper sulfate solution.

[0058] In some embodiments of the present invention, in step S2, a Lix84 extractant having a volume content of 10% to 15% is used.

[0059] In some embodiments of the present invention, in step S2, the acidity of sulfuric acid is 4N.

[0060] In step S2 of the present invention, the Ni and Cu in the leaching solution are extracted into the organic phase using the Lix84 extractant. Since the Lix84 extractant does not extract Co 3+ and Zn 2+, it is left in the raffinate to be separated from Ni and Cu; then 4N sulfuric acid is used to strip Ni in the organic phase, and the pH of the stripped liquid is controlled to be 1.5-2.0 to obtain nickel sulfate liquid, which can be directly used for ternary batching, or crystallized into nickel sulfate crystals or electrolytically deposited metallic nickel products; the organic phase after stripping Ni is stripped of Cu with 4N sulfuric acid, and the acidity of the stripped liquid is controlled to be 2N to obtain copper sulfate liquid, which is sent to the copper electrolytic deposition process to recover copper.

[0061] In some embodiments of the present invention, in step S3, the pH of the raffinate is greater than 8.

[0062] In some embodiments of the present invention, in step S3, the temperature of the ammonia evaporation treatment is 85-95°C; typically but not limitatively, for example, in step S3, the temperature of the ammonia evaporation treatment can be 85°C, 90°C, 95°C or a range of any two thereof.

[0063] In some embodiments of the present invention, in step S3, the raffinate is subjected to ammonia evaporation to obtain ammonia gas, cobalt-zinc slag and residual night; the evaporated ammonia gas is absorbed by water and the obtained ammonia water is returned to step S1 for ammonia complex leaching; the residual liquid is returned to step S1 for ammonia complex leaching.

[0064] In step S3 of the present invention, since the pH of the raffinate is greater than 8, it can be directly subjected to ammonia evaporation treatment, and the evaporated ammonia gas can be absorbed by water to form ammonia water, which can be returned to the ammonia complex leaching process for use; after the ammonia is evaporated, Co 3+ and Zn 2+ The complex ions are lost and become free ions. In an alkaline environment, they become hydroxides and precipitate to form cobalt-zinc slag. The final residual liquid returns to the ammonia complex leaching process and is reused as bottom water.

[0065] In some embodiments of the present invention, in step S4, acid leaching includes: adding sulfuric acid solution to cobalt-zinc slag, and performing acid leaching at a pH of 1 to 1.5 to obtain a zinc-containing solution and cobalt slag; the zinc-containing solution is a zinc sulfate solution; typically but not limitatively, for example, the pH of the acid leaching can be 1, 1.1, 1.2, 1.3, 1.4, 1.5 or a range of any two thereof.

[0066] In some embodiments of the present invention, in step S4, the acid leaching temperature is 20-50°C; the acid leaching time is 4-6h; typically but not limitatively, for example, in step S4, the acid leaching temperature can be 20°C, 30°C, 40°C, 50°C or a range consisting of any two thereof; the acid leaching time can be 4h, 5h, 6h or a range consisting of any two thereof.

[0067] In step S4 of the present invention, the cobalt-zinc slag is subjected to dilute acid leaching, and sulfuric acid leaching is used in an acidic environment with a pH of 1 to 1.5. When the pH value remains unchanged, the leaching is performed to obtain zinc sulfate solution, and the Co 3+ It is not dissolved under the above acidity and remains in the slag to form cobalt slag; the zinc sulfate solution obtained by leaching can be crystallized into zinc sulfate crystals, or it can be cast into zinc ingots after electrolytic deposition of metallic zinc.

[0068] In some embodiments of the present invention, in step S5, reduction leaching comprises: adding a sulfuric acid solution and a reducing agent to the cobalt slag, and performing reduction leaching at a pH of 0.5 to 1.5 to obtain a cobalt-containing solution; the cobalt-containing solution is a cobalt sulfate solution; typically but not limitatively, for example, the pH of the reduction leaching is 0.5, 0.8, 1, 1.2, 1.5 or a range of any two thereof.

[0069] In some embodiments of the present invention, in step S5, the reducing agent includes at least one of hydrogen peroxide, sodium sulfite and sodium metabisulfite.

[0070] In some embodiments of the present invention, in step S5, the amount of reducing agent used is 1.2 to 1.5 times the theoretical amount; the theoretical amount refers to the amount of reducing agent required for the chemical reaction of reducing and leaching cobalt, and different reducing agents have different theoretical amounts.

[0071] In some embodiments of the present invention, in step S5, the temperature of the reduction acid leaching is 40-60°C; the time of the reduction acid leaching is 4-6 hours; typically but not limitatively, for example, the temperature of the reduction acid leaching can be 40°C, 50°C, 60°C or a range consisting of any two thereof; the time of the reduction acid leaching can be 4h, 5h, 6h or a range consisting of any two thereof.

[0072] In step S5 of the present invention, a reducing agent is added to the cobalt slag in an acidic environment with a pH of 1±0.5, and sulfuric acid is used for leaching. After complete dissolution, a cobalt sulfate liquid is obtained; the cobalt sulfate liquid can be directly used for ternary batching, and can also be used to crystallize cobalt sulfate crystals or electrolytically deposit metallic cobalt products.

[0073] The method for recovering metals from crude nickel and cobalt hydroxide of the present invention can obtain five major products, namely nickel sulfate, cobalt sulfate, zinc sulfate, copper sulfate and manganese slag, without the discharge of waste water and waste slag, and can achieve zero emission.

[0074] Example 1

[0075] The method for recovering metal from crude nickel cobalt hydroxide provided in this embodiment comprises the following steps:

[0076] S1. After adding water to crude nickel cobalt hydroxide (MHP) to prepare the slurry, a slurry with a solid-liquid ratio of 1:3 is obtained; 70% of the weight of MHP ammonium bicarbonate and 4.6 times the weight of MHP ammonia water with a concentration of 25% are added to the slurry, leaching at 50°C for 4 hours, and then air is introduced at 50°C for oxidation for 4 hours to obtain a leachate and manganese slag; wherein the composition of the crude nickel cobalt hydroxide is as shown in Table 1; the composition of the leachate is as shown in Table 2; the pH of the leachate is 9.4; the composition of the manganese slag is as shown in Table 3; the leaching rate of cobalt is 96.44%, the leaching rate of nickel is 99.11%, the leaching rate of copper is 99.45%, and the leaching rate of zinc is 96.67%.

[0077] Table 1

[0078] element Co Ni Cu Fe Ca Mg Mn Zn Content (wt%) 4.08 39.17 0.14 0.031 0.086 0.9 7.26 0.85

[0079] Table 2

[0080] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 1.91 20.88 0.075 0.0007 0.0005 0.0005 0.0005 0.35

[0081] Table 3

[0082] element Co Ni Cu Mn Zn Content (wt%) 0.73 1.93 0.012 32.5 0.12

[0083] S2. The leachate is extracted with a Lix84 extractant having a volume content of 15% to obtain an extracted organic phase and a raffinate; the extracted organic phase is subjected to a first back extraction with 4N sulfuric acid to obtain a nickel sulfate solution and a first organic phase; the first organic phase is subjected to a second back extraction with 4N sulfuric acid to obtain a copper sulfate solution; wherein the composition of the raffinate is as shown in Table 4, and the pH of the raffinate is 8.7; the composition of the nickel sulfate solution is as shown in Table 5, and the pH of the nickel sulfate solution is 1.8; the composition of the copper sulfate solution is as shown in Table 6, and the acidity of the copper sulfate solution is 2.1N; the cobalt recovery rate is 99.5%, the nickel recovery rate is 99.5%, the copper recovery rate is 99.5%, and the zinc recovery rate is 99.7%.

[0084] Table 4

[0085] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 1.47 0.0001 0.0001 0.0005 0.0003 0.0004 0.0005 0.30

[0086] Table 5

[0087] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.019 125.09 0.0001 0.0002 0.0005 0.0003 0.0005 0.0003

[0088] Table 6

[0089] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.0002 0.001 63.12 0.0001 0.0003 0.0002 0.0002 0.0002

[0090] S3, treating the raffinate with ammonia at 90°C to obtain ammonia gas, cobalt-zinc slag and residual night; the ammonia water obtained by absorbing the evaporated ammonia gas with water is returned to step S1 for ammonia complex leaching; the residual liquid is returned to step S1 as a bottom liquid for ammonia complex leaching; wherein the composition of the cobalt-zinc slag is shown in Table 7; since the residual liquid after ammonia evaporation is returned to ammonia complex leaching, the recovery rates of cobalt and zinc are both 100%.

[0091] Table 7

[0092] element Co Ni Cu Fe Ca Mg Mn Zn Content (wt%) 38.54 0.001 0.0005 0.0002 0.0004 0.0005 0.0005 17.61

[0093] S4. Add sulfuric acid solution to the cobalt-zinc slag, and carry out acid leaching for 4 hours at 45° C. and pH 1.4 to obtain zinc sulfate solution and cobalt slag; the composition of the zinc sulfate solution is shown in Table 8, and the pH of the zinc sulfate solution is 1.4; the recovery rate of zinc is 99.5%.

[0094] Table 8

[0095] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.001 0.0005 0.0028 0.001 0.003 0.0025 0.0031 102.32

[0096] S5. Add sulfuric acid solution to the cobalt slag, maintain the pH at 1.4 at 45°C, slowly add 1.4 times the theoretical amount of hydrogen peroxide, the total leaching time is 6 hours, and obtain cobalt sulfate solution; the composition of the cobalt sulfate solution is shown in Table 9, and the pH of the cobalt sulfate solution is 1.4; the recovery rate of cobalt is 99.9%.

[0097] Table 9

[0098] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 110.23 0.001 0.0001 0.0002 0.0002 0.0003 0.0004 0.001

[0099] Example 2

[0100] The method for recovering metal from crude nickel cobalt hydroxide provided in this embodiment comprises the following steps:

[0101] S1. After adding water to crude nickel cobalt hydroxide (MHP) to prepare a slurry, a slurry with a solid-liquid ratio of 1:3 is obtained; 70% of the weight of MHP ammonium bicarbonate and 4.6 times the weight of MHP ammonia water with a concentration of 25% are added to the slurry, leaching at 38° C. for 6 hours, and then air is introduced at 38° C. for oxidation for 3.5 hours to obtain a leachate and manganese slag; wherein the composition of the crude nickel cobalt hydroxide is the same as that in Example 1; the composition of the leachate is shown in Table 10; the pH of the leachate is 8.7; the composition of the manganese slag is shown in Table 11; the leaching rate of cobalt is 96.48%, the leaching rate of nickel is 99.21%, the leaching rate of copper is 99.66%, and the leaching rate of zinc is 97%.

[0102] Table 10

[0103] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 2.01 21.74 0.08 0.0006 0.0004 0.0005 0.0005 0.4

[0104] Table 11

[0105] element Co Ni Cu Mn Zn Content (wt%) 0.62 1.56 0.01 35.8 0.11

[0106] S2. The leachate was extracted with a Lix84 extractant having a volume content of 12% to obtain an extracted organic phase and a raffinate; the extracted organic phase was subjected to a first back extraction with 4N sulfuric acid to obtain a nickel sulfate solution and a first organic phase; the first organic phase was subjected to a second back extraction with 4N sulfuric acid to obtain a copper sulfate solution; wherein the composition of the raffinate is as shown in Table 12, and the pH of the raffinate is 8.3; the composition of the nickel sulfate solution is as shown in Table 13, and the pH of the nickel sulfate solution is 1.5; the composition of the copper sulfate solution is as shown in Table 14, and the acidity of the copper sulfate solution is 2.0N; the cobalt recovery rate is 99.6%, the nickel recovery rate is 99.5%, the copper recovery rate is 99.5%, and the zinc recovery rate is 99.6%.

[0107] Table 12

[0108] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 1.65 0.0001 0.0001 0.0004 0.0003 0.0003 0.0004 0.35

[0109] Table 13

[0110] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.017 121.35 0.0001 0.0002 0.0005 0.0003 0.0005 0.0003

[0111] Table 14

[0112] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.0001 0.001 64.57 0.0001 0.0002 0.0002 0.0002 0.0001

[0113] S3, subjecting the raffinate to ammonia evaporation at 85°C to obtain ammonia gas, cobalt-zinc slag and residual night; the evaporated ammonia gas is absorbed by water to obtain ammonia water, which is returned to step S1 for ammonia complex leaching; the residual liquid is returned to step S1 as a bottom liquid for ammonia complex leaching; wherein the composition of the cobalt-zinc slag is as shown in Table 15; since the residual liquid after ammonia evaporation is returned to ammonia complex leaching, the recovery rates of cobalt and zinc are both 100%.

[0114] Table 15

[0115] element Co Ni Cu Fe Ca Mg Mn Zn Content (wt%) 38.79 0.0004 0.0004 0.0002 0.0004 0.0003 0.0003 18.23

[0116] S4. Add sulfuric acid solution to the cobalt-zinc slag, and carry out acid leaching for 5.5 hours at 25° C. and pH 1.1 to obtain zinc sulfate solution and cobalt slag; the composition of the zinc sulfate solution is shown in Table 16, and the pH of the zinc sulfate solution is 1.1; the recovery rate of zinc is 99.5%.

[0117] Table 16

[0118] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.0008 0.0003 0.0024 0.0009 0.002 0.0015 0.0011 100.54

[0119] S5. Add sulfuric acid solution to the cobalt slag, maintain the pH at 1.3 at 55°C, slowly add 1.2 times the theoretical amount of sodium sulfite, and the total leaching time is 5 hours to obtain cobalt sulfate solution; the composition of the cobalt sulfate solution is shown in Table 17, and the pH of the cobalt sulfate solution is 1.3; the recovery rate of cobalt is 99.9%.

[0120] Table 17

[0121] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 110.23 0.001 0.0001 0.0002 0.0002 0.0003 0.0004 0.001

[0122] Example 3

[0123] The method for recovering metal from crude nickel cobalt hydroxide provided in this embodiment comprises the following steps:

[0124] S1. After adding water to crude nickel cobalt hydroxide (MHP) to prepare the slurry, a slurry with a solid-liquid ratio of 1:3 is obtained; 70% of the weight of MHP ammonium bicarbonate and 4.6 times the weight of MHP ammonia water with a concentration of 25% are added to the slurry, leaching at 45°C for 5 hours, and then air is introduced at 45°C for oxidation for 3 hours to obtain a leachate and manganese slag; wherein the composition of the crude nickel cobalt hydroxide is the same as that in Example 1; the composition of the leachate is shown in Table 18; the pH of the leachate is 10.1; the composition of the manganese slag is shown in Table 19; the leaching rate of cobalt is 96.22%, the leaching rate of nickel is 99.14%, the leaching rate of copper is 99.57%, and the leaching rate of zinc is 96.87%.

[0125] Table 18

[0126] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 1.98 21.34 0.078 0.0004 0.0003 0.0003 0.0004 0.37

[0127] Table 19

[0128] element Co Ni Cu Mn Zn Content (wt%) 0.65 1.43 0.01 35.5 0.12

[0129] S2. The leachate is extracted with 10% by volume of Lix84 extractant to obtain an extracted organic phase and a raffinate; the extracted organic phase is subjected to a first back extraction with 4N sulfuric acid to obtain a nickel sulfate solution and a first organic phase; the first organic phase is subjected to a second back extraction with 4N sulfuric acid to obtain a copper sulfate solution; wherein, the composition of the raffinate is as shown in Table 20, and the pH of the raffinate is 9.2; the composition of the nickel sulfate solution is as shown in Table 21, and the pH of the nickel sulfate solution is 2.0; the composition of the copper sulfate solution is as shown in Table 22, and the acidity of the copper sulfate solution is 1.9N; the cobalt recovery rate is 99.5%, the nickel recovery rate is 99.6%, the copper recovery rate is 99.6%, and the zinc recovery rate is 99.7%.

[0130] Table 20

[0131] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 1.88 0.0001 0.0001 0.0003 0.0002 0.0002 0.0003 0.35

[0132] Table 21

[0133] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.014 120.35 0.0001 0.0001 0.0003 0.0002 0.0005 0.0003

[0134] Table 22

[0135] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.0001 0.001 65.32 0.0001 0.0002 0.0002 0.0002 0.0001

[0136] S3, subjecting the raffinate to ammonia evaporation at 95°C to obtain ammonia gas, cobalt-zinc slag and residual night; the evaporated ammonia gas is absorbed by water to obtain ammonia water, which is returned to step S1 for ammonia complex leaching; the residual liquid is returned to step S1 as a bottom liquid for ammonia complex leaching; wherein the composition of the cobalt-zinc slag is as shown in Table 23; since the residual liquid after ammonia evaporation is returned to ammonia complex leaching, the recovery rates of cobalt and zinc are both 100%.

[0137] Table 23

[0138] element Co Ni Cu Fe Ca Mg Mn Zn Content (wt%) 39.12 0.0005 0.0004 0.0002 0.0003 0.0003 0.0004 18.79

[0139] S4. Add sulfuric acid solution to the cobalt-zinc slag, and carry out acid leaching for 4.5 hours at 35° C. and pH 1.25 to obtain zinc sulfate solution and cobalt slag; the composition of the zinc sulfate solution is shown in Table 24, and the pH of the zinc sulfate solution is 1.25; the recovery rate of zinc is 99.5%.

[0140] Table 24

[0141] element Co Ni Cu Fe Ca Mg Mn Zn Content (g / L) 0.0009 0.0006 0.0021 0.0008 0.0015 0.0025 0.001 101.74

[0142] S5. Add sulfuric acid solution to the cobalt slag, maintain the pH at 1.2 at 50°C, slowly add 1.3 times the theoretical amount of sodium metabisulfite, the total leaching time is 4 hours, and obtain cobalt sulfate solution; the composition of the cobalt sulfate solution is shown in Table 25, the pH of the cobalt sulfate solution is 1.2; the recovery rate of cobalt is 99.9%.

[0143] Table 25

[0144]

[0145]

[0146] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention, rather than to limit it. Although the present invention has been described in detail with reference to the aforementioned embodiments, those skilled in the art should understand that they can still modify the technical solutions described in the aforementioned embodiments, or replace some or all of the technical features therein with equivalents. However, these modifications or replacements do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of the present invention.

Claims

1. A method for recovering metals from crude nickel and cobalt hydroxide, characterized in that: The steps include: S1, subjecting the crude nickel and cobalt hydroxide to ammonia complex leaching and oxidation in sequence to obtain a leachate and manganese slag; S2, the leachate is separated by extraction to obtain a raffinate, a nickel-containing solution and a copper-containing solution; S3, subjecting the raffinate to ammonia evaporation treatment to obtain cobalt-zinc slag; S4, acid leaching the cobalt-zinc slag to obtain a zinc-containing solution and cobalt slag; S5, reducing and leaching the cobalt slag to obtain a cobalt-containing solution.

2. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 1, characterized in that: In step S1, the ammonia complex leaching comprises: adding ammonium salt and ammonia water to the slurry obtained after slurrying of crude nickel cobalt hydroxide, and leaching at below 50° C. for 4 to 6 hours.

3. The method for recovering metal from crude nickel and cobalt hydroxide according to claim 2, characterized in that: In step S1, the ammonium salt includes ammonium bicarbonate and / or ammonium carbonate.

4. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 2, characterized in that: In step S1, the mass ratio of the crude nickel cobalt hydroxide to the ammonium salt is 100:(65-80); And / or, the usage ratio of the crude nickel cobalt hydroxide and the ammonia water is 100g:400-600mL.

5. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 2, characterized in that: In step S1, the oxidation includes: introducing air into the slurry after the ammonia complex leaching for oxidation.

6. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 5, characterized in that: In step S2, the extraction separation comprises: extracting the leaching solution with Lix84 extractant to obtain an extracted organic phase and a raffinate; and subjecting the extracted organic phase to secondary stripping to obtain the nickel-containing solution and the copper-containing solution.

7. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 6, characterized in that: In step S2, the extracted organic phase is subjected to a first back extraction with sulfuric acid to obtain the nickel-containing solution and the first organic phase, and the pH of the nickel-containing solution is 1.5-2.0; the first organic phase is subjected to a second back extraction with sulfuric acid to obtain the copper-containing solution, and the acidity of the copper-containing solution is 1.8-2.2N.

8. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 7, characterized in that: In step S4, the acid leaching includes: adding a sulfuric acid solution to the cobalt-zinc slag, and performing acid leaching at a pH of 1 to 1.5 to obtain the zinc-containing solution and the cobalt slag.

9. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 8, characterized in that: In step S5, the reduction leaching comprises: adding a sulfuric acid solution and a reducing agent to the cobalt slag, and performing reduction leaching at a pH of 0.5 to 1.5 to obtain the cobalt-containing solution.

10. The method for recovering metals from crude nickel and cobalt hydroxide according to claim 9, characterized in that: In step S5, the reducing agent includes at least one of hydrogen peroxide, sodium sulfite and sodium metabisulfite.

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