A process for comprehensively recovering valuable metals from cobalt-rich waste materials

Abstract

The present invention relates to a process for comprehensively recovering valuable metals from cobalt-rich waste. Cobalt silicate in the waste is converted into sodium cobalt silicate through sodium carbonate roasting pretreatment, cobalt ferrite is converted into cobaltous oxide and ferric oxide, and composite oxides of tungsten, tantalum, and niobium are converted into soluble sodium salts such as sodium tungstate and sodium chromate. These are then separated from alkaline oxides such as cobaltous oxide in a subsequent water leaching process. Cobalt, iron, and copper in the water leaching residue are leached into a solution through sulfuric acid reduction leaching, while metals such as tantalum and niobium are contained in the acid leaching residue. Tungsten and chromium in the water leaching solution are separated by concentrated hydrochloric acid neutralization followed by tungsten precipitation followed by chromium reduction precipitation. This process has the advantages of a high elemental leaching rate and high separation efficiency of cobalt from elements such as tungsten and chromium. Furthermore, the process is simple, the reaction conditions are mild, and the raw materials used are inexpensive. It can achieve comprehensive recovery of elements such as cobalt, tungsten, and tantalum.

Description

Technical Field

[0001] The present invention belongs to the technical field of nonferrous metal smelting, and specifically relates to a process for comprehensively recovering valuable metals from cobalt-rich waste, and in particular to a process for comprehensively recovering valuable metals such as cobalt, tungsten, chromium, and tantalum from cobalt-rich waste. Background Art

[0002] With the development of the non-ferrous metallurgical industry, the comprehensive utilization of secondary metallurgical resources, mainly smelting waste slag and urban mines, has received increasing attention. Recycling the valuable metals therein can not only save resources but also avoid environmental pollution. The material involved in this patent is cobalt-rich waste slag. Cobalt is an important raw material and is widely used in various fields such as chemical industry, medicine, aerospace, etc. From the perspective of both resource conservation and environmental protection, it is very necessary to extract cobalt and other valuable metals from this cobalt-rich waste material to achieve resource utilization and harmless treatment.

[0003] There are numerous methods for separating cobalt from cobalt-containing raw materials. The methods employed vary depending on the cobalt's presence in different raw materials. For example, hydrocobaltite, primarily composed of CoOOH, copper-cobalt oxide ores primarily composed of CoOOH, CuCo2S4, and CoCuCO3, and lithium-ion battery cathode materials primarily composed of LiCoO2 are often treated with reduction acid leaching. Cobalt oxide ores primarily composed of m(Co,Ni)O·MnO2·nH2O can also be recovered by reduction smelting to produce cobalt-iron alloy phases and slag. Converter slag, primarily composed of cobalt silicate (Co2SiO4) and cobalt ferrite (CoFe2O4) with low cobalt grades, is treated by pyrometallurgical enrichment followed by hydrometallurgical separation of cobalt, nickel, and copper. Laterite nickel ore, where cobalt is embedded in goethite or hematite, can be treated with ammonium salt roasting followed by ammonia leaching, reduction roasting followed by ammonia leaching, or sulfuric acid roasting. Hydrometallurgical zinc smelting often involves direct acid leaching after roasting to purify cobalt slag. Cobalt white alloys, scrap cemented carbide, and scrap high-temperature alloys containing cobalt in elemental form can be recovered using electrochemical dissolution or oxidation followed by acid leaching. The primary phases of these materials are cobalt silicate (Co2SiO4) and composite oxides of cobalt, iron, and copper (CoFe2O4 and Co3O4). There are few reports in the literature on the recovery of cobalt-containing materials primarily containing these phases, and the commonly used reduction acid leaching method is ineffective. Furthermore, these waste materials contain elements such as tungsten and chromium, which are less well-recovered by existing processes.

[0004] Applying existing treatment processes for cobalt-containing materials to this material results in low leaching rates and separation efficiencies. Furthermore, the material has a complex elemental composition, containing valuable elements such as tungsten, tantalum, and niobium, as well as harmful elements such as chromium. Summary of the Invention

[0005] The present invention provides a process for comprehensively recovering valuable metals from cobalt-rich waste. The technical problem to be solved is to solve the problems of low leaching rate and low efficiency in separating cobalt from other metals in existing processes for this material.

[0006] In order to solve the above technical problems, the present invention provides a process for comprehensively recovering valuable metals from cobalt-rich waste, which is characterized by comprising the following steps:

[0007] Step 1: The cobalt-rich waste material that has been dried and ground and sieved to a size of less than 0.074 mm is thoroughly mixed with sodium carbonate, and the mixture is heated to 700-900° C. in an air atmosphere for roasting, whereby the acidic or amphoteric oxide reacts with the sodium carbonate to form a corresponding sodium salt and the cobalt, iron, and copper composite oxides decompose to obtain roasted clinker;

[0008] Step 2: adding deionized water to the roasted clinker obtained in step 1 to dissolve the soluble sodium salt therein, and filtering to obtain a water leaching solution containing tungsten, chromium, aluminum, and silicon and a water leaching residue enriched with cobalt, copper, tantalum, and niobium;

[0009] Step 3: Leaching the water leaching residue obtained in step 2 with sulfuric acid to recover the cobalt and copper elements therein, adding a reducing agent to leach the cobalt and copper to obtain a leaching residue rich in titanium, tantalum and niobium;

[0010] Step 4: The pH of the aqueous extract obtained in step 2 is adjusted with dilute hydrochloric acid, and the dilute hydrochloric acid is added dropwise while stirring to separate the silicon and aluminum elements therein to obtain a desiliconized liquid and silicon slag I;

[0011] Step 5: boiling and concentrating the desiliconized liquid obtained in step 4 and filtering to obtain a desiliconized concentrated liquid and silicon slag II;

[0012] Step 6: Add the desiliconized concentrated solution obtained in step 5 to concentrated hydrochloric acid, stir, and filter to obtain tungstic acid and tungsten precipitation solution;

[0013] Step 7: adding sodium sulfite solution to the tungsten precipitation solution for reduction, adding sodium hydroxide solution to neutralize the precipitation, and filtering to obtain chromium residue and chromium precipitation solution.

[0014] Beneficial Effects: The present invention uses sodium carbonate roasting pretreatment to convert the cobalt silicate in the waste into sodium cobalt silicate, cobalt ferrite into cobaltous oxide and ferric oxide, and composite oxides of tungsten, tantalum, and niobium into soluble sodium salts such as sodium tungstate and sodium chromate, thereby achieving efficient separation of cobalt from tungsten and chromium, and further achieving comprehensive recovery of valuable metal elements. The present invention is characterized by high leaching rates for metals such as cobalt, tungsten, and copper. The present invention has efficient separation and does not produce excess waste. It enriches tungsten into tungstic acid and chromium into trivalent chromium slag. The acid leaching slag contains high-grade elements such as tantalum, niobium, and titanium, which facilitates further recovery. The present invention has mild reaction conditions and is easy to implement, with potential for industrialization. BRIEF DESCRIPTION OF THE DRAWINGS

[0015] Figure 1 It is a process flow chart of the present invention;

[0016] Figure 2 This is a process flow chart for recovering tungsten and chromium. DETAILED DESCRIPTION

[0017] In order to make the purpose, content and advantages of the present invention more clear, the specific embodiments of the present invention are further described in detail below.

[0018] The present invention proposes a process for comprehensively recovering valuable metals from cobalt-rich waste, comprising the following steps:

[0019] Step 1: The cobalt-rich waste that has been dried and ground and sieved to a size below 0.074 mm is thoroughly mixed with sodium carbonate. The mixture is heated to about 900°C at a certain heating rate in an air atmosphere to react with acidic or amphoteric oxides such as tungsten and chromium with sodium carbonate to form corresponding sodium salts, and the cobalt, iron, and copper composite oxides are decomposed:

[0020] WO3+Na2CO3=Na2WO4+CO2↑

[0021] CrO3+Na2CO3=Na2CrO4+CO2↑

[0022] CoFe2O4=CoO+Fe2O3Cu 0.76 Co 2.24 O4+0.76CoO→0.76CuO+Co3O4

[0023] Co2SiO4+Na2CO3→Na2CoSiO4+CoO+CO2↑

[0024] Step 2: Add the roasted clinker obtained in step 1 into deionized water to dissolve the soluble sodium salt therein, and filter to obtain a water leaching solution containing tungsten, chromium, aluminum, and silicon and a water leaching residue enriched with metals such as cobalt, copper, tantalum, and niobium.

[0025] Step 3: Leach the water-leached residue obtained in step 2 with sulfuric acid to recover elements such as cobalt and copper, and add sodium sulfite as a reducing agent. The reaction needs to be carried out under closed conditions.

[0026] After leaching cobalt and copper, the leaching residue rich in titanium, tantalum and niobium is obtained, and the reaction occurs:

[0027] Co(Fe, Cu)O+H2SO4→Co(Fe, Cu)SO4+H2O

[0028] Fe2O3+2H2SO4+Na2SO3→2FeSO4+Na2SO4+2H2O

[0029] Co3O4+Na2SO3+3H2SO4→3CoSO4+Na2SO4+3H2O

[0030] Na2CoSiO4+H2SO4→CoSO4+Na2SiO3+H2O

[0031] Step 4: Adjust the pH of the water extract obtained in step 2 with dilute hydrochloric acid, adding dilute hydrochloric acid while stirring. This step can separate silicon, aluminum and other elements, and obtain silicon-free liquid and silicon slag, which react:

[0032] 2xNaAlO2+yNa2SiO3+(x+y-1)H2O→

[0033] Na2O·xAl2O3·ySiO2↓+2(x+y-1)NaOH

[0034] Step 5: Boil and concentrate the desiliconized liquid obtained in step 4 and filter to obtain desiliconized concentrated liquid and silicon slag.

[0035] Step 6: Add the desiliconized concentrated solution obtained in step 5 to concentrated hydrochloric acid, stir, and filter to obtain tungstic acid and tungsten precipitation solution. Sodium chromate reacts to form sodium dichromate, and the reaction occurs:

[0036] Na2WO4+2HCl(conc.)→2NaCl+H2WO4↓

[0037] 2Na2CrO4+2HCl→Na2Cr2O7+2NaCl

[0038] Step 7: Add sodium sulfite solution to the tungsten precipitation solution for reduction, add sodium hydroxide solution to neutralize the precipitation, and filter to obtain chromium residue and chromium precipitation solution, and react:

[0039] Na2WO4+2HCl(cone.)→2NaCl+H2WO4↓

[0040] 2Na2CrO4+2HCl→Na2Cr2O7+2NaCl

[0041] The components of the cobalt-rich waste in step 1 of the present invention include cobalt silicate Co2SiO4, a composite oxide of cobalt and copper Cu 2.24 Co 0.76 O4, composite oxide of cobalt and iron CoFe2O4, composite oxide of tungsten, tantalum and niobium Ta(Nb)8W9O 47 . Its main metal element composition is: Co 20-35%, Fe 8-15%, Cu 8-15%, Ni 0-3%, W 0-3%, Cr 0-3%, Al 0-3%, Ti 0-3%, Ta 0-1%, Nb 0-1%.

[0042] In the present invention, the amount of sodium carbonate added in step 1 is 0.7-1.5 times the mass of the cobalt-rich waste according to the standard that acidic oxides such as tungsten and chromium in the cobalt-rich waste are converted into corresponding sodium salts. The roasting atmosphere is air, the heating rate is 0.01-10°C / min, and the roasting time is 1-2h.

[0043] In the present invention, the liquid-to-solid ratio during the water immersion in step 2 is 3-5 mL / g, and washing is required after leaching, with a washing liquid-to-solid ratio of 1-3 mL / g and a washing time of 20 min. The water immersion reaction is carried out at room temperature.

[0044] The reducing agent used in the reduction acid leaching described in step 3 of the present invention is sulfur dioxide or sodium sulfite, sodium thiosulfate, or the like, which reacts with acid to generate sulfur dioxide gas. The sulfuric acid concentration is 1-1.5 mol / L, the acid leaching temperature is 90-95°C, and the acid leaching time is 20-60 minutes. If sodium sulfite is used as the reducing agent, the amount of sodium sulfite added is 30% of the mass of the leaching residue (dry).

[0045] The concentration of dilute hydrochloric acid used in step 4 of the present invention is 1-4 mol / L, the stirring rate is 50-200 r / min, the pH of the water extract is adjusted to 7-12, and the solution is allowed to stand and filtered after reacting for 30 minutes.

[0046] In step 5 of the present invention, the desiliconized solution is boiled and concentrated, which can promote the further occurrence of the desiliconization reaction and also promote the further concentration of the solution until the concentrations of sodium tungstate and sodium chromate are both 10-25 g / L.

[0047] The mixing method in step six of the present invention is to add the concentrated solution into concentrated hydrochloric acid, and the volume of the concentrated hydrochloric acid is 0.3-1.0 times the volume of the concentrated solution.

[0048] In step seven, 1-1.5 mol / L sodium sulfite solution is added and the potential of the solution is adjusted to below 450 mV. The pH of the solution is adjusted to 9-11 with sodium hydroxide solution. The solution is stirred for 10-60 minutes and then filtered.

[0049] Example 1 is the result of treatment according to the process flow of the present invention, and Examples 2 to 4 are the results of treatment with other processes, which are used for comparative experiments.

[0050] Example 1:

[0051] (1) Sodium carbonate roasting pretreatment

[0052] Cobalt-rich waste, dried, ground, and sieved to a size of less than 200 mesh, contained the following elements: Co 26.73%, Cu 8.90%, Fe 6.79%, Cr 2.26%, Ti 1.77%, Ni 1.52%, W 2.29%, Al 2.06%, Ta 0.70%, and Nb 0.42%. Sodium carbonate (20.0 g of the original waste and 15.0 g of sodium carbonate) was added at 0.75 times the mass of the waste and mixed thoroughly. The mixture was then heated to 900°C in a muffle furnace at a heating rate of 8°C / min. After calcination at 900°C for 2 hours, the mixture was cooled to below 100°C and removed. A yield of 22.1 g of clinker was obtained, yielding approximately 63.1%.

[0053] (2) Flooding

[0054] The roasted clinker is crushed and ground to a size of less than 200 mesh. Deionized water at a liquid-to-solid ratio of 3:1 is then mixed with the clinker and stirred at room temperature for a water leaching reaction. After leaching, the suspension is filtered through a 0.45 μm solvent filter and washed with deionized water at a liquid-to-solid ratio of 3:1. This produces a water leaching solution primarily composed of sodium tungstate and sodium chromate, and a water leaching residue primarily composed of cobalt, iron, and copper oxides, tantalum, and sodium niobate. The leaching rate for tungsten reaches 99.9%, the leaching rate for chromium reaches 99.5%, and the leaching rate for cobalt is close to zero. 100 g of clinker yields 62.4 g of water leaching residue and 575 mL of water leaching solution. The element contents in the water leaching solution are approximately: 2.33 g / L tungsten; 2.25 g / L chromium; 0.42 g / L aluminum; and 0.02 g / L tantalum.

[0055] (3) Acid leaching

[0056] Dry the leaching residue in an oven at 80°C for more than 24 hours, grind it in an agate mortar, and sieve it to less than 0.075mm (-200 mesh). Place the leaching residue and sodium sulfite powder (30% of the leaching residue's mass) in a beaker, add 1.5mol / L dilute sulfuric acid at a liquid-to-solid ratio of 12:1, and quickly seal the beaker to prevent the escape of SO2. Heat the reaction in a 95°C water bath for 30 minutes. After the reaction, filter the suspension through a solvent filter with a pore size of 0.45μm. Centrifugal filtration before membrane filtration can increase the filtration rate. Under these optimized conditions, 10g of dried leached residue was used to produce 132mL of acid leaching solution (the residue was filtered and washed with a small amount of deionized water, resulting in a larger volume than 120mL) and 0.196g of leached residue. The element contents in the leaching solution were 22.2g / L Co, 7.65g / L Cu, 5.43g / L Fe, 1.14g / L Ti, 0.56g / L Ta, and 0.29g / L Nb. The leaching efficiency was 99.97% for cobalt, 99.89% for copper, 100% for iron (no iron was detected after dissolution of the leached residue), 77.0% for titanium, 83.5% for tantalum, and 72.3% for niobium.

[0057] (4) Separation of tungsten and chromium from aqueous solution

[0058] The aqueous extract was adjusted to a pH of 9-11 with 2 mol / L dilute hydrochloric acid to produce sodium silicate slag to remove impurities such as silicon and aluminum. The silicon removal rate reached 97.2%, and the resulting sodium silicate slag contained 21.19% silicon, 17.58% aluminum, and 10.79% sodium. The desiliconized solution was boiled, concentrated, and filtered. The filtrate (i.e., the tungsten / sodium chromate concentrate) was cooled to room temperature and added to 0.6 volumes of concentrated hydrochloric acid. The mixture was stirred at 100 rpm for 12 hours, allowed to stand, and then filtered. The tungsten precipitation rate was 99.6%, and the resulting tungstic acid contained 77.58% tungsten and zero chromium. The tungsten precipitation solution was then adjusted to a potential of 450 mV with 1.5 mol / L sodium sulfite solution and a pH of 9 with sodium hydroxide solution. The chromium slag and chromium precipitation solution were then filtered. The chromium content in the chromium slag was approximately 37.03%.

[0059] Example 2:

[0060] Take 5g of cobalt-rich waste that was dried, ground and sieved to below 200 mesh, which contained the following elements: Co 26.73%, Cu 8.90%, Fe 6.79%, Cr 2.26%, Ti 1.77%, Ni 1.52%, W 2.29%, Al 2.06%, Ta 0.70%, Nb 0.42%. 500mL of 2mol / L H2SO4 was added at a liquid-solid ratio of 10:1, and leached at 70°C for 4h. The leaching rate of cobalt was 87.3%, the leaching rate of copper was 70.2%, the leaching rate of tantalum was 10.4%, the leaching rate of tungsten was 3.7%, and the residue rate was 27.0%.

[0061] Example 3:

[0062] Take 5g of cobalt-rich waste that is dried, ground and sieved to below 200 mesh, which contains the following elements: Co 26.73%, Cu 8.90%, Fe 6.79%, Cr 2.26%, Ti 1.77%, Ni 1.52%, W 2.29%, Al 2.06%, Ta 0.70%, Nb 0.42%. Add 500mL 2mol / L H2SO4 at a liquid-solid ratio of 10:1, and leaching at 20°C for 24h. The leaching rate of cobalt is 83.4%, the leaching rate of copper is 70.2%, the leaching rate of tantalum is 10.4%, the leaching rate of tungsten is 3.7%, and the residue rate is 22.2%.

[0063] Example 4:

[0064] Take 5g of cobalt-rich waste that is dried, ground and sieved to below 200 mesh, which contains the following elements: Co 26.73%, Cu 8.90%, Fe 6.79%, Cr 2.26%, Ti 1.77%, Ni 1.52%, W 2.29%, Al 2.06%, Ta 0.70%, Nb 0.42%. Add 500mL 2mol / L H2SO4 at a liquid-solid ratio of 10:1, and add 30% H2O2 at a liquid-solid ratio of 5:1. Leach at 70°C for 24h. The leaching rate of cobalt is 72.6%, the leaching rate of copper is 63.7%, the leaching rate of tantalum is 32.3%, the leaching rate of tungsten is 1.7%, and the residue rate is 26.2%.

[0065] The present invention uses sodium carbonate roasting pretreatment to convert the cobalt silicate in the waste into sodium cobalt silicate, the cobalt ferrite into cobaltous oxide and ferric oxide, and the composite oxides of tungsten, tantalum, and niobium into soluble sodium salts such as sodium tungstate and sodium chromate. These are then separated from alkaline oxides such as cobaltous oxide in a subsequent water leaching process. Cobalt, iron, and copper in the water leaching residue are leached into a solution through sulfuric acid reduction leaching, while metals such as tantalum and niobium are present in the acid leaching residue. Tungsten and chromium in the water leaching solution are separated by neutralization with concentrated hydrochloric acid to precipitate tungsten and then reduction to precipitate chromium. This process has the advantages of a high elemental leaching rate and high separation efficiency of cobalt from elements such as tungsten and chromium. It also features a simple process, mild reaction conditions, and inexpensive raw materials, enabling comprehensive recovery of elements such as cobalt, tungsten, and tantalum.

[0066] The above is only a preferred embodiment of the present invention. It should be pointed out that for ordinary technicians in this technical field, several improvements and modifications can be made without departing from the technical principles of the present invention. These improvements and modifications should also be regarded as the scope of protection of the present invention.

Claims

1. A process for comprehensive recovery of valuable metals from cobalt-rich waste, characterized in that: The following steps are involved: Step 1: The cobalt-rich waste material that has been dried and ground and sieved to a size of less than 0.074 mm is thoroughly mixed with sodium carbonate, and the mixture is heated to 700-900° C. in an air atmosphere for roasting, whereby the acidic or amphoteric oxide reacts with the sodium carbonate to form a corresponding sodium salt and the cobalt, iron, and copper composite oxides decompose to obtain roasted clinker; Cobalt-rich waste contains cobalt, copper, iron, tungsten, chromium, tantalum and niobium; the components of cobalt-rich waste include cobalt silicate Co2SiO4, cobalt ferrite CoFe2O4, and composite oxide of cobalt and copper Cu 2.24 Co 0.76 O4, cobalt tetraoxide Co3O4, composite oxide of tungsten, tantalum and niobium Ta(Nb)8W9O 47 ; Step 2: adding deionized water to the roasted clinker obtained in step 1 to dissolve the soluble sodium salt therein, and filtering to obtain a water leaching solution containing tungsten, chromium, aluminum, and silicon and a water leaching residue enriched with cobalt, copper, tantalum, and niobium; Step 3: Leaching the water leaching residue obtained in step 2 with sulfuric acid to recover the cobalt and copper elements therein, adding a reducing agent to leach the cobalt and copper to obtain a leaching residue rich in titanium, tantalum and niobium; Step 4: The pH of the aqueous extract obtained in step 2 is adjusted with dilute hydrochloric acid, and the dilute hydrochloric acid is added dropwise while stirring to separate the silicon and aluminum elements therein to obtain a desiliconized liquid and silicon slag I; Step 5: boiling and concentrating the desiliconized liquid obtained in step 4 and filtering to obtain a desiliconized concentrated liquid and silicon slag II; Step 6: Add the desiliconized concentrated solution obtained in step 5 to concentrated hydrochloric acid, stir, and filter to obtain tungstic acid and tungsten precipitation solution; Step 7: adding sodium sulfite solution to the tungsten precipitation solution for reduction, adding sodium hydroxide solution to neutralize the precipitation, and filtering to obtain chromium residue and chromium precipitation solution.

2. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The main metal element composition of cobalt-rich waste is: Co 20-35%, Fe 8-15%, Cu 8-15%, Ni 0-3%, W 0-3%, Cr0-3%, Al 0-3%, Ti 0-3%, Ta 0-1%, Nb 0-1%.

3. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The amount of sodium carbonate added in step 1 is 0.7-1.5 times the mass of the cobalt-rich waste according to the standard of converting the acidic oxides of tungsten and chromium in the cobalt-rich waste into corresponding sodium salts respectively; the roasting temperature in step 1 is 900°C, the roasting time is 1-3h, the heating rate during roasting is 5-10°C / min, and the atmosphere during roasting is an oxidizing atmosphere.

4. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The second step of water immersion is carried out at room temperature, with a liquid-to-solid ratio of 3-10 mL / g and a immersion time of 10-20 min. Water immersion requires washing, with a washing liquid-to-solid ratio of 1-3 mL / g.

5. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The reducing agent used in the sulfuric acid leaching in step 3 is sulfur dioxide or sodium sulfite or sodium thiosulfate which can react with acid to generate sulfur dioxide gas.

6. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: In step 3, the sulfuric acid concentration during sulfuric acid leaching is 1-3 mol / L, the leaching temperature during sulfuric acid leaching is 70-120° C., the leaching time during sulfuric acid leaching is 20-60 min, and the sodium sulfite added in step 3 is 30% of the mass of the water leaching residue.

7. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The concentration of the dilute hydrochloric acid used in step 4 is 1-4 mol / L. In step 4, the pH of the water extract is adjusted to 7-12. Stirring is required during pH adjustment in step 4. The stirring rate is 50-200 r / min and the stirring time is 10-60 min.

8. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: In step 5, the desiliconized solution is boiled and concentrated to promote further desiliconization reaction and further concentration of the solution until the concentrations of sodium tungstate and sodium chromate are both 10-25 g / L.

9. A process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: The mixing method in step six is to add the concentrate into concentrated hydrochloric acid, the volume of concentrated hydrochloric acid is 0.3-1.0 times the volume of the concentrate, the mixing temperature is room temperature 20-30° C., and the reaction time in step six is 0.5-12 h.

10. The process for comprehensive recovery of valuable metals from cobalt-rich waste according to claim 1, characterized in that: In step seven, the concentration of the added sodium sulfite solution is 1-1.5 mol / L, and the sodium sulfite solution is added to adjust the potential of the solution to below 450 mV. In step seven, the pH of the solution is adjusted to 9-11 with sodium hydroxide solution. Stirring is required during pH adjustment, the stirring rate is 20-100 r / min, and the stirring time is 10-60 min.

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