CN111778404A - A kind of leaching separation method of nickel-cobalt-molybdenum-phosphorus-vanadium alloy material - Google Patents

Abstract

The invention discloses a method for leaching and separating nickel-cobalt-molybdenum-phosphorus-vanadium alloy materials. According to the method of the invention, the molybdenum, phosphorus and vanadium in the alloy material are reacted with alkali by slurrying the ingredients to generate corresponding molybdate, phosphate and vanadate ions in the form of entering the solution, and then Pressurized oxidative alkali leaching, through the control of the pressure parameters in the reaction kettle, the amount and rate of the introduction of oxygen, and the control of the reaction temperature in the reaction kettle, the molybdenum, phosphorus, vanadium and alkali in the feed liquid are controlled. The reaction is all converted into sodium molybdate, sodium phosphate and sodium vanadate, and nickel, cobalt and iron enter the slag in the form of precipitation, so as to ensure that the valuable metals including molybdenum, phosphorus and vanadium in the feed liquid can fully enter the filtrate for supply. Subsequent separation and recycling to achieve full recycling of valuable metal resources in the alloy material.

Description

A kind of leaching separation method of nickel-cobalt-molybdenum-phosphorus-vanadium alloy material

technical field

The invention relates to the technical field of metallurgy and chemical industry, in particular to a method for leaching and separating nickel-cobalt-molybdenum-phosphorus-vanadium alloy materials.

Background technique

At present, nickel-cobalt alloy is a strategic metal with ferromagnetic properties, which is widely used in the production of stainless steel, superalloys, magnetic materials and catalysts, and its application fields cover petrochemical, aerospace, military, electronics and other fields. The amount of alloy waste such as spent catalysts and nickel-cobalt superalloys is also increasing. The above-mentioned wastes are not only large in amount, but also contain relatively high content of rare metals, including valuable metals such as nickel (Ni), cobalt (Co), molybdenum (Mo), phosphorus (P), and vanadium (V). If it is disposed of at will, it will not only cause serious pollution hazards to the environment, but also waste most of the rare metal resources. Recycling it as a secondary resource can not only directly obtain certain economic benefits, but also improve the utilization rate of resources, avoid environmental problems caused by catalysts, and achieve sustainable development.

In order to fully recycle the rare valuable metal resources in the above-mentioned nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy materials, the common recycling method at present usually adopts acid leaching at normal pressure, adding sulfuric acid, hydrochloric acid or nitric acid and an oxidizing agent, under high temperature and high acid conditions. Next, various metal elements in the waste catalyst are leached into the solution, and then the method of step-by-step extraction is used to extract and purify molybdenum, cobalt and nickel respectively, and then various metals are recovered by evaporation and crystallization. This method is quicker for leaching, but requires high acid, high temperature, and the acid concentration reaches 2 mol/L before cobalt and nickel can be leached. Moreover, in this case, the nickel, cobalt, iron, and molybdenum in the leaching solution are relatively high, and the acidity is also relatively high. In the process of adding oxidant to remove iron, iron is oxidized to trivalent, and it is easy to react with molybdenum and vanadium to form iron molybdate and iron vanadate into the slag, resulting in low recovery rate of molybdenum and vanadium valuable metals. In the subsequent extraction process, since molybdenum exists in the form of molybdate radicals, while cobalt and nickel exist in the form of cations, the extraction process control is very complicated, which is difficult for industrial production; the final metal recovery rate is low, generally only reach 80% to 85%.

The Chinese invention patent with the application number CN109517988A discloses a method for leaching and separating cobalt-nickel alloy materials containing molybdenum and vanadium. The metals such as cobalt, nickel, molybdenum and vanadium are separated from iron slag by means of pressurized acid leaching and oxygen enrichment, effectively avoiding the The oxidized ferric iron reacts with molybdovanadate ions, resulting in the loss of molybdenum and vanadium metal elements. Finally, molybdenum and vanadium are leached from nickel and cobalt in an acid solution step by step. Through the above method, the recovery rate of molybdenum and vanadium is increased to more than 97%, but the technical defect of this method is: separation of molybdenum and vanadium in the acid cobalt-nickel solution requires multiple extractions, thereby causing a part of molybdenum to contain vanadium, and a part of molybdenum to contain vanadium. Vanadium contains nickel, and further separation is difficult, resulting in a low recovery rate of the final molybdenum vanadium.

SUMMARY OF THE INVENTION

The object of the present invention is to overcome the low recovery rate of molybdenum and vanadium in the cobalt-nickel alloy containing molybdenum and vanadium existing in the prior art, the process operation is more complicated, the time consuming is longer, and the process cost is high and insufficient, and a kind of nickel-cobalt-molybdenum is provided. A method for leaching and separating phosphorus-vanadium alloy materials.

In order to achieve the above-mentioned purpose of the invention, the present invention provides the following technical solutions:

A method for leaching and separating alloy materials, comprising the steps of:

S1, raw material processing: the alloy waste is made into alloy powder; the alloy waste contains iron, cobalt, nickel, molybdenum, phosphorus and vanadium elements;

S2, batching slurry reaction: the alloy powder and the alkaline solution are prepared into a material liquid according to the solid-liquid mass ratio of 1:7-15, and then stirred at 70-90 ℃, and the slurry reaction is 0.5-1h; the alkaline solution The mass ratio of liquid caustic soda or flake caustic soda and alloy powder in the solution is 1.2-3.5:1;

S3. Pressurized oxidative alkali leaching: pump the reacted liquid in S2 into the reaction kettle, inject oxygen to increase the pressure to 1.8-2.8MPa, and react at 160-250 ° C for 2-8h; adjust the slurry after the reaction is completed pH is 7-9, filter, obtain filter residue and filtrate, described filter residue contains iron, cobalt, nickel element, and described filtrate contains molybdate ion, phosphate ion and vanadate ion;

S4, multi-metal separation: the iron, cobalt and nickel elements in the S3 filter residue are separated, and the molybdenum, phosphorus and vanadium elements in the S3 filtrate are separated.

In the above-mentioned processing method, the feed liquid and the alkaline solution react as follows:

MoO ₃ +2NaOH==Na ₂ MoO ₄ +H ₂ O

V ₂ O ₅ +2NaOH==2NaVO ₃ +H ₂ O

P ₂ O ₅ +6NaOH==2Na ₃ PO ₄ +3H ₂ O

According to the leaching and separation method of the present invention, by subjecting the alloy material containing nickel, cobalt, molybdenum, phosphorus and vanadium to the slurry reaction of the ingredients, the molybdenum, phosphorus and vanadium in the alloy material are reacted with alkali to generate corresponding molybdate and phosphoric acid. After the root and vanadate ion forms enter the solution, the feed liquid after the above-mentioned batching slurry reaction is pumped into the reaction kettle. Through the control of the pressure parameters in the reaction kettle, the amount of oxygen introduced and the rate of introduction are combined with The control of the reaction temperature in the reaction kettle makes the reaction of molybdenum, phosphorus, vanadium and alkali in the feed liquid all convert into sodium molybdate, sodium phosphate and sodium vanadate, and nickel, cobalt and iron enter the slag in the form of precipitation. Ensure that the valuable metals including molybdenum, phosphorus and vanadium in the feed liquid can fully enter the filtrate for subsequent separation and recycling, so as to fully recycle the valuable metal resources in the alloy material. Molybdenum, phosphorus, vanadium and other metals are separated in the alkaline filtrate, the process is simple, the time-consuming is short, and the purity of each metal after separation is good.

In the alloy material, the weight percent content of each element is:

The content of molybdenum is 9-30%; the content of phosphorus is 5-20%; the content of vanadium is 5-20%; the content of nickel is 10-40%; the content of cobalt is 5-30%;

It has been proved by experimental research that the content of molybdenum, phosphorus and vanadium valuable metals in the filter residue obtained after the batching slurry reaction and the pressure leaching reaction of the present invention can be controlled at a maximum of 0.1% or less, so that the valuable metals in the alloy feed liquid can be controlled. Almost all of it enters the filtrate for later separation and recycling.

As a preferred version of the present invention, the separation of molybdenum, phosphorus and vanadium elements in the S4 filtrate comprises the following steps:

S411, low-temperature phosphorus removal: the mixed filtrate in S3 is subjected to low-temperature crystallization, water washing, and filtration to obtain sodium phosphate crystals and the second filtrate, and the low-temperature crystallization temperature is 1-5 ° C;

S412a, separation of molybdenum: adjusting the pH of the filtrate obtained in S411 to 7.0-9.0, and at 25-60° C., the second filtrate is adsorbed by a molybdenum-absorbing resin to obtain a filtrate after removing molybdenum ions; the molybdenum-absorbing resin is to be adsorbed After saturation, decompose to obtain molybdate solution;

S413a, separating vanadium: the filtrate after removing molybdenum ions is adsorbed by a vanadium-absorbing resin, and after the vanadium-absorbing resin is saturated with adsorption, it is analyzed to obtain a vanadate solution.

Wherein, the speed of the filtrate of S411 passing through the molybdenum-absorbing resin is 5-10 m ³ /h; the speed of the filtrate after removing molybdenum ions in the S412 passing through the vanadium-absorbing resin is 5-10 m ³ /h.

The above separated molybdenum and vanadium metals are recovered by precipitation to obtain molybdenum and vanadium products.

The molybdate solution separated by the molybdenum-absorbing resin in step S412a is precipitated and desiliconized by the flocculant magnesium sulfate, and then adjusted to pH 1-2 with hydrochloric acid, and industrial-grade molybdenum is precipitated at 30-50° C. acid products;

The vanadate solution separated by the vanadium-absorbing resin in step S413a is precipitated and desiliconized by the flocculant aluminum sulfate, and the pH of the solution is adjusted to about 1.5-3.5 at 70-80° C., and ammonium chloride is used to precipitate an industrial-grade solution. of ammonium metavanadate products.

As a preferred solution of the present invention, the analysis solution used in the analysis process in the step S412a and the step S413a is a sodium hydroxide solution with a mass fraction of 10-20%.

As a preferred solution of the present invention, the model of the molybdenum-absorbing resin is any one of ZGD314, D352 and PDM.

As a preferred solution of the present invention, the model of the vanadium-absorbing resin is ZGD231 or LS-32.

After analysis, molybdenum-absorbing resin and vanadium-absorbing resin are washed and regenerated with sulfuric acid solution with a mass fraction of 5-20% and pure water, and can be recycled until the pH of the effluent reaches 2.0-5.0.

In order to further optimize the amount of solvent, the inventors found that the filtrate after phosphorus removal was precipitation of vanadium element using ammonium sulfate solution to directly obtain amine metavanadate, and then the solution after removal of vanadium was reacted with perchloric acid solution to obtain trioxide Molybdenum precipitation. In this way, it is avoided that after the resin adsorbs vanadium and molybdenum elements, excess alkaline solution is needed to resolve the vanadium and molybdenum elements from the resin, and then the vanadium-absorbing resin and molybdenum-absorbing resin need to be recovered by an acidic solution. Return it to its original state for recycling. The use of acidic and alkaline solutions is further saved. further reducing costs. Therefore, it is further preferred that the second filtrate obtained in S411 is separated in the following manner:

S412b, the preparation of technical grade ammonium metavanadate: use the aqueous solution of ammonium sulfate whose mass fraction is 25%-35%, adjust the pH of the second filtrate to 7.0～9.5, at 20～75 ℃, stir, filter , obtain technical grade ammonium metavanadate and the fourth filtrate containing sodium molybdate;

Preparation of S413b and molybdenum trioxide: adjust the pH of the fourth filtrate containing sodium molybdate obtained from S412 to 0.5-4.5 with perchloric acid whose mass fraction is 60%-70%, and stir at 30-80°C, Filter and dry to obtain molybdenum trioxide.

Among them, in the process of separating molybdenum and vanadium, the chemical reaction equation is:

2NaVO ₃ +(NH ₄ ) ₂ SO ₄ == 2NH ₄ VO ₃ ↓+Na ₂ SO ₄

Na ₂ MoO ₄ +H ₂ O+2HClO ₄ ==MoO ₃ (H ₂ O) ₂ +2NaClO ₄

MoO ₃ (H ₂ O) ₂ ==MoO ₃ +2H ₂ O

As a preferred version of the present invention, described S4, the separation of iron, cobalt and nickel elements in the filter residue comprises the following steps:

S421, pressurized oxidative acid leaching: the filter residue containing iron, cobalt and nickel elements in the S3 and the acidic solution are prepared into a feed liquid according to the solid-liquid mass ratio of 1:6-15, stirred at 70-90 ° C, and slurried After reacting for 1-5h, pump it into a reactor and pressurize it to 1-2.8Mpa, and after reacting for 2-8h, filter to obtain iron slag and filtrate, and the filtrate contains nickel ions and cobalt ions;

S422, extracting and separating nickel and cobalt elements, and sequentially chemically removing impurities and extracting the filtrate in S421 to obtain a cobalt sulfate solution and a nickel sulfate solution; the extraction agent is P507;

The mass ratio of sulfuric acid and the slag containing iron, cobalt and nickel metals in the acidic solution is 1.5-3.5:1.

Chemical impurity removal is to remove trace impurities such as Zn, Mn, Cu, Fe, Ca and so on in the solution through P204 extraction, and the extracted nickel sulfate and cobalt sulfate solutions are evaporated, concentrated and crystallized to recover battery-grade cobalt sulfate and battery-grade nickel sulfate products. .

As a preferred solution of the present invention, in the S421, the pH of the reacted feed solution is adjusted to 2.0-4.0, and the iron slag and the filtrate containing nickel ions and cobalt ions are obtained by filtration.

As a preferred solution of the present invention, reverse washing is performed at least twice before the iron, cobalt and nickel metals in the S4 slag are separated, until the pH of the filtrate during the washing process is 7.0-8.0.

Among them, the "reverse washing" method specifically refers to the use of clean water for the second washing, and the second washing water is used for the first washing of the new slag, and the first washing water is reused in step S1. The circulating reverse washing method used by the liquid. According to the above-mentioned 2 reverse washing methods, and the pH of the washing solution in the overall circulating washing process is controlled to be 8.0 to 14.0 by means of alkali adjustment. On the one hand, the content of molybdenum, phosphorus and vanadium valuable metals in the alloy material can be further leached. , so that the valuable metals enter the solution and then return to the batching slurry reaction step to fully separate, leaching and recover. On the other hand, it can effectively avoid the environmental pollution problem caused by the discharge of washing waste liquid, and make the process operation more environmentally friendly.

As a preferred solution of the present invention, the particle size of the alloy powder is 100-200 mesh.

Pre-fine grinding and sieving to a particle size of about 100 mesh can increase the specific surface area of the alloy material, increase the contact area of the subsequent reaction, and be more conducive to the full progress of the reaction.

Compared with the prior art, the beneficial effects of the present invention:

1. According to the method of the present invention, after the cobalt-nickel alloy material containing molybdenum and vanadium is subjected to the step S2 batching slurry reaction, in the step S3 adding alkali and pressure leaching molybdenum, phosphorus and vanadium, the reaction kettle The optimal control of each reaction parameter in the feed solution makes the molybdenum, phosphorus and vanadate ions in the feed solution react with the alkali, so as to ensure that the valuable metals of molybdenum, phosphorus and vanadium in the feed solution can fully enter the filtrate for subsequent separation and recycling, and realize the alloy Full recovery and utilization of valuable metal resources in the material.

2. According to the scheme of the present invention, through low-temperature washing to remove phosphorus, a mixed solution of high-purity solid sodium phosphate, sodium molybdate and sodium vanadate is obtained, and the recovery rate of phosphorus is more than 98.5%.

3. According to the scheme of the present invention, molybdenum and vanadium metals are separated by resin, and after the adsorption of the molybdenum-absorbing resin and vanadium-absorbing resin is saturated, the molybdate solution and vanadate solution are obtained respectively by analyzing the molybdenum-absorbing resin and vanadium-absorbing resin, respectively. The overall recovery of vanadium was >98.5%.

4. According to the scheme of the present invention, the total recovery rate of nickel and cobalt can reach >99%, and the valuable metals are basically completely recovered.

5. According to the solution of the present invention, the separation process of molybdenum, phosphorus and vanadium metals in an alkaline solution is simpler, the separation is more thorough, and the purity of the separated metals is high.

6. In the process of extracting molybdenum element and vanadium element by chemical precipitation method, vanadium is precipitated by ammonium sulfate to form industrial grade ammonium metavanadate, molybdenum is precipitated by perchloric acid and dried to form molybdenum trioxide, and the recovery rate of molybdenum and vanadium is high. High purity, reducing the process flow of resin column, can reduce the amount of acid and alkali.

Description of drawings:

Fig. 1 is the process flow diagram of the present invention;

Fig. 2 is another process flow diagram of the present invention.

Detailed ways

The present invention will be further described in detail below in conjunction with test examples and specific embodiments. However, it should not be construed that the scope of the above-mentioned subject matter of the present invention is limited to the following embodiments, and all technologies realized based on the content of the present invention belong to the scope of the present invention.

Example 1

As shown in Figure 1, take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after electric furnace mixed reduction and smelting, take 1kg material as an example, obtain nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material through electric furnace mixed reduction and smelting, after testing the alloy material middle:

The content of Co is 9.36%, the content of Ni is 30.15%, the content of Mo is 15.65%, the content of Fe is 14.23%, the content of V is 9.4%, and the content of P is 12.35%; the above alloy material is processed as follows:

The alloy material is finely sieved to about 100 meshes, and water and liquid caustic soda are added to the alloy material according to the solid-liquid mass ratio of 1:10 and the alkali material mass ratio of 2:1, respectively, to obtain the slurry after the slurry, The reaction was then stirred at 75°C for 3 hours.

Pressurized oxidative alkali leaching: pump the slurry after the slurry reaction into the pressurized reaction kettle, after sealing the reaction kettle, turn on the steam heating, introduce oxygen, and increase the pressure in the kettle to 1.2MPa, then slow down the oxygen injection Speed, slowly increase the pressure to 1.9MPa within 60 minutes, during the pressure increase process, after the temperature rises to 180℃, close the steam valve, start timing for 2 hours, and control the pressure between 1.90 and 2.0Mpa during the heat preservation process. , open the steam valve to heat to maintain the temperature between 220 and 250 ° C, and keep the temperature for 2 hours. After the insulation is completed, close the steam valve, close the oxygen valve, then slowly open the emptying valve, and open the cooling water of the reactor. After the internal pressure is reduced to normal pressure, and the temperature in the kettle is reduced to 80 °C, the slurry in the reactor is pumped into the normal pressure reaction tank, and filtered to obtain nickel-cobalt-containing iron slag and sodium molybdate, sodium phosphate, and sodium vanadate. mixture.

The mixed solution of sodium molybdate, sodium phosphate and sodium vanadate obtained above is subjected to low-temperature dephosphorization, and the obtained mixed solution containing sodium molybdate, sodium phosphate and sodium vanadate is washed with low-temperature water, and the temperature of low-temperature water is controlled at about 2 ℃, and then filtered A mixed solution of solid high-purity sodium phosphate, sodium molybdate and sodium vanadate is obtained.

After testing, the content of each metal in the leaching solution obtained after filtration is: Mo: 13.45g/L, P: 12.87g/L, V: 8.5g/L, and the slag rate is about 80%. The content of each metal in the solid sodium phosphate obtained after washing with low temperature water is: Mo: 0.0023%, V: 0.0045%, P: 19.28%, which can be sent to a phosphate fertilizer plant for recycling.

Get the mixed solution of above-mentioned filtered sodium molybdate and sodium vanadate, adjust pH to about 8.0, and control at 40 ° C temperature, pass through ZGD324 molybdenum-absorbing resin with a flow rate of 7m per hour, until ^ZGD324 absorbs molybdenum-absorbing resin adsorption saturation . After testing, the content of each metal in the solution adsorbed by ZGD324 molybdenum-absorbing resin is: Mo: 0.002g/L, V: 8.06g/L.

The above-mentioned solution after being adsorbed by ZGD324 molybdenum-absorbing resin is controlled to pass through ZGD201 vanadium-absorbing resin at a flow rate of 8m3 until ZGD201 vanadium-absorbing resin is saturated with adsorption. After testing, the content of each metal in the solution after adsorption and saturation by ZGD201 vanadium absorbing resin is: Mo: 0.0001g/L, V: 0.001g/L.

After the resin is saturated, 10% sodium hydroxide solution is used as the desorption solution to backwash the saturated ZGD324 molybdenum-absorbing resin and ZGD324 molybdenum-absorbing resin to obtain sodium molybdate solution and sodium vanadate solution, respectively.

After testing, the molybdenum content in the sodium molybdate solution obtained by analysis was 80g/L; the vanadium content in the sodium vanadate solution obtained by analysis was 64g/L.

Further, the analyzed resin was washed with 6% sulfuric acid solution to wash and regenerate the resin, and then washed with pure water until the pH of the effluent stopped at 7.0, and the resin could be recycled for use.

Further, the alkali leaching alloy material is continuously added with acid and pressurized and oxidized, and the obtained nickel-cobalt-containing iron slag is added to the alloy material according to the solid-liquid mass ratio of 1:10 and the acid-material mass ratio of 2.5:1, respectively, to add water and sulfuric acid, After the slurry was obtained, the slurry was stirred and reacted for 4 hours at 70° C. The slurry was pumped into the pressurized reaction kettle, steam heating was turned on, and oxygen was introduced to make the pressure in the kettle rise to 1.0 MPa. , slow down the oxygen supply speed, slowly increase the pressure to 2.8MPa, and control the temperature in the reactor to 230 ° C, after 3 hours of heat preservation reaction, open the emptying valve to reduce the pressure in the reactor to normal pressure, and use soda ash to call back After the pH reaches 2.5, the recovered iron slag and the nickel-cobalt-containing filtrate are obtained by filtration. After filtration, the above-mentioned nickel-cobalt-containing filtrate was taken, and the Co content in the liquid was 8.69 g/L, the Ni content was 26.35 g/L, and the Fe content was 4.75 g/L.

Extraction and separation of nickel and cobalt: The solution containing nickel and cobalt is first extracted with P204 to remove trace impurities such as Zn, Mn, Cu, Fe, Ca, etc. in the solution, and then the extraction and separation coefficients of nickel and cobalt are different with P507, and the solution is extracted in turn. The nickel and cobalt in the solution are stripped with sulfuric acid to obtain a high-purity nickel sulfate solution and a high-purity cobalt sulfate solution, respectively. Further, the extracted organic phase can be returned for repeated use.

The nickel sulfate and cobalt sulfate products are recovered by evaporation, concentration and crystallization, and the cobalt sulfate solution and the nickel sulfate solution obtained above are respectively subjected to evaporation concentration and crystallization to obtain battery-grade cobalt sulfate and battery-grade nickel sulfate products. After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 99.12%; nickel recovery rate: 99.27%; molybdenum recovery rate: 99.06%; phosphorus recovery rate: 98.58%; vanadium recovery rate: 98.93%.

Example 2

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 10.01%, the content of Ni is 30.9%, the content of Mo is 13.52%, the content of Fe is 13.51%, and the content of V is 8.48%; the above alloy materials are processed as follows:

The specific technique is the same as that described in Example 1. The only difference is that:

Pressurized oxidative alkali leaching: pump the slurry after the slurry reaction into the pressurized reaction kettle, after sealing the reaction kettle, turn on steam heating, introduce oxygen, and increase the pressure in the kettle to 1.6MPa, then slow down the oxygen injection Speed, slowly increase the pressure to 2.5MPa within 60 minutes, in the process of pressure increase, after the temperature rises to 190 ℃, close the steam valve, start timing for 1 hour, and control the pressure between 2.5-2.6Mpa during the heat preservation process , open the steam valve to heat to maintain the temperature between 190-230 ° C, and keep it for 1 hour. After the insulation is over, close the steam valve, close the oxygen valve, then slowly open the emptying valve, open the cooling water of the reactor, and put the cooling water in the reactor. After the pressure is reduced to normal pressure and the temperature in the kettle is reduced to 80 °C, the slurry in the reactor is pumped into the normal pressure reaction tank, and filtered to obtain a mixture of nickel-cobalt-containing iron slag and sodium molybdate, sodium phosphate and sodium vanadate. solution.

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 99.25%; nickel recovery rate: 99.30%; molybdenum recovery rate: 99.27%; phosphorus recovery rate: 98.62%; vanadium recovery rate: 99.14%.

Example 3

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 9.82%, the content of Ni is 31.55%, the content of Mo is 14.86%, the content of Fe is 13.38%, and the content of V is 8.76%; the above alloy materials are processed as follows:

The specific technique is the same as that described in Example 1. The only difference is that:

Pressurized oxidative alkali leaching: pump the slurry after the slurry reaction into the pressurized reaction kettle, after sealing the reaction kettle, turn on the steam heating, introduce oxygen, and increase the pressure in the kettle to 1.2MPa, then slow down the oxygen injection Speed, slowly increase the pressure to 1.7MPa within 60 minutes, during the pressure increase process, after the temperature rises to 180℃, close the steam valve, start timing for 2.5 hours, and control the pressure between 1.7-1.9Mpa during the heat preservation process , open the steam valve to heat to maintain the temperature between 180-220 ° C, and keep it for 2.5 hours. After the insulation is over, close the steam valve, close the oxygen valve, then slowly open the emptying valve, open the cooling water of the reactor, and put the cooling water in the reactor. After the pressure is reduced to normal pressure and the temperature in the kettle is reduced to 80 °C, the slurry in the reactor is pumped into the normal pressure reaction tank, and filtered to obtain a mixture of nickel-cobalt-containing iron slag and sodium molybdate, sodium phosphate and sodium vanadate. solution.

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 99.39%; nickel recovery rate: 99.58%; molybdenum recovery rate: 99.81%; phosphorus recovery rate: 99.68%; vanadium recovery rate: 99.18%.

Example 4

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 9.25%, the content of Ni is 30.58%, the content of Mo is 15.82%, the content of Fe is 14.37%, and the content of V is 9.06%; the above alloy materials are processed as follows:

The specific process technique is shown in Figure 2, and the difference with the process described in Embodiment 1 is only:

Molybdenum and vanadium were extracted separately: the frozen molybdenum and vanadium solution after dephosphorization was used an aqueous solution of ammonium sulfate, the mass fraction was 30%, the pH was adjusted to 8.5, stirred at 70 ° C for 2 hours, and filtered to obtain technical grade ammonium metavanadate and containing For the filtrate of sodium molybdate, adjust the pH of the filtrate containing sodium molybdate to 0.5-4.5 with 70% perchloric acid, stir at 75° C. for 2 hours, filter and dry to obtain molybdenum trioxide.

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 99.38%; nickel recovery rate: 99.59%; molybdenum recovery rate: 99.90%; phosphorus recovery rate: 99.67%; vanadium recovery rate: 99.33%.

Example 5

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 10.06%, the content of Ni is 31.02%, the content of Mo is 13.94%, the content of Fe is 14.68%, and the content of V is 8.95%; the above alloy materials are processed as follows:

The specific process technique is shown in Figure 2, and the difference with the process described in Embodiment 2 is only:

Molybdenum and vanadium were extracted separately: the frozen molybdenum and vanadium solution after dephosphorization was used an aqueous solution of ammonium sulfate, the mass fraction was 30%, the pH was adjusted to 7.0-9.5, stirred at 25 ° C for 2 hours, and filtered to obtain technical grade ammonium metavanadate and sodium molybdate-containing filtrate, adjust the pH of the sodium molybdate-containing filtrate to 0.5-4.5 with 60% perchloric acid, stir at 40° C. for 2 h, filter and dry to obtain molybdenum trioxide.

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 99.41%; nickel recovery rate: 99.63%; molybdenum recovery rate: 99.92%; phosphorus recovery rate: 99.77%; vanadium recovery rate: 99.43%.

Comparative Example 1

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 9.98%, the content of Ni is 29.84%, the content of Mo is 12.58%, the content of Fe is 13.63%, and the content of V is 7.76%; the above alloy materials are processed as follows:

The specific technique is the same as that described in Example 1. The only difference is:

Pressurized oxidative alkali leaching: pump the slurry after the slurry reaction into the pressurized reaction kettle, after sealing the reaction kettle, turn on the steam heating, introduce oxygen, and increase the pressure in the kettle to 1.0MPa, then slow down the oxygen injection Speed, slowly increase the pressure to 1.5MPa within 60 minutes, in the process of pressure increase, after the temperature rises to 180 ℃, close the steam valve, start timing and keep warm for 2.5 hours, and control the pressure between 1.5-1.7Mpa during the heating process , open the steam valve to heat to maintain the temperature between 160-190 ° C, and keep it for 2.5 hours. After the insulation is completed, close the steam valve, close the oxygen valve, then slowly open the emptying valve, and open the cooling water of the reactor. After the pressure is reduced to normal pressure and the temperature in the kettle is lowered to 80 °C, the slurry in the reactor is pumped into the normal pressure reaction tank, and filtered to obtain a mixture of nickel-cobalt-containing iron slag and sodium molybdate, sodium phosphate and sodium vanadate. solution.

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 88.12%; nickel recovery rate: 73.28%; molybdenum recovery rate: 78.36%; phosphorus recovery rate: 82.94%; vanadium recovery rate: 70.48%.

Comparative Example 2

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 9.67%, the content of Ni is 28.64%, the content of Mo is 14.46%, the content of Fe is 13.41%, the content of P is 11.95%, and the content of V is 9.52%; the above alloy materials are processed as follows:

The specific technique is the same as that described in Example 1. The only difference is that in the process of separating iron, cobalt and nickel, the parameters in the process of acid pressure oxidation are adjusted:

Adding acid and pressurizing oxidation to the filter residue of the alkali leaching alloy material, adding water and sulfuric acid to the alloy material according to the solid-liquid mass ratio of 1:10 and the acid-material mass ratio of 2.5:1 to obtain the obtained nickel-cobalt-containing iron slag to obtain After slurrying, the feed liquid was stirred and reacted at 70°C for 4 hours, the feed liquid after the slurry reaction was pumped into the pressurized reaction kettle, the steam heating was turned on, and oxygen was introduced to make the pressure in the kettle rise to 0.5MPa, Slow down the oxygen supply speed, make the pressure rise slowly to 1.2MPa, and control the temperature in the reactor to 150 °C. After holding the reaction for 3 hours, open the vent valve to reduce the pressure in the reactor to normal pressure, and adjust the pH with soda ash. = 2.5, and then filtered to obtain recovered iron slag and nickel-cobalt-containing filtrate. After filtration, the above-mentioned nickel-cobalt-containing filtrate was taken, and the Co content in the liquid was 5.31 g/L, the Ni content was 11.29 g/L, and the Fe content was 3.68 g/L.

Comparative Example 3

Take the nickel-cobalt-molybdenum-phosphorus-vanadium-containing alloy material obtained after mixing reduction and smelting in an electric furnace, and take 1kg material as an example, after testing the alloy material:

The content of Co is 10.67%, the content of Ni is 29.65%, the content of Mo is 13.46%, the content of Fe is 14.51%, the content of P is 11.35%, and the content of V is 9.12%; the above alloy materials are processed as follows:

The specific technique is the same as that described in Example 1. The only difference is:

This process is to carry out pressurized oxygen-enriched acid leaching first:

(1) grinding: the alloy material is ground and sieved to below 100 mesh, and then the alloy powder after grinding and sieving is processed as follows:

(2) Batching slurry reaction: according to the liquid-solid ratio of 10:1 and the acid-to-material ratio of 2:1 to the above-mentioned alloy powder, after adding water and sulfuric acid to mix, react at 75 ° C for 3 hours to obtain a slurry reaction post-feed.

(3) Iron removal by pressurized leaching: then pump the slurry after the slurry reaction into the pressurized reaction kettle, after sealing the reaction kettle, turn on steam heating, introduce oxygen, and adjust the pressure to 1.2Mpa. Then slowly increase the pressure from 1.2Mpa to 1.9Mpa within 60 minutes, and close the steam valve after the temperature rises to 180°C during the pressure increase process. Start timing and keep warm for 2 hours, control the pressure between 1.90 and 2.0Mpa and the temperature between 200 and 230 ℃ during the heat preservation process. After another 2 hours of heat preservation, when the heat preservation is over, close the steam valve, close the oxygen valve, and then slowly Open the emptying valve, open the cooling water of the reaction kettle, after the pressure in the reaction kettle drops to normal pressure and the temperature in the kettle drops to 80 °C, the slurry in the reaction kettle is pumped into the normal pressure reaction tank, and the The slurry is pumped into the atmospheric pressure reaction tank, soda ash is slowly added, pH is adjusted to 1.5, and liquid-solid separation is performed to obtain recovered iron slag and filtrate. The above-mentioned gained reclaims iron slag, and then through 2 reverse washings (the second washing water removes the first washing, the first washing water returns to the above-mentioned batching slurry reaction and is used as a dosing solution, and the second washing adds clear water), washing During the process, the pH of the washing solution is controlled to be 3, and the iron slag is recovered after washing is obtained. After testing, the content of each metal in the leaching solution obtained after filtration is: Co: 11.86g/L, Ni: 19.26g/L, Mo: 17.69g/L, V: 11.15g/L. The content of each metal in the recovered iron slag obtained after washing is: Co: 0.03%, Ni: 0.035%, Mo: 1.093%, P: 1.106%; V: 1.075%; the slag rate is 22.1%, and the iron content is 43.7% .

After the whole treatment process, through analysis and calculation, the weight percent content of each element is:

Cobalt recovery rate: 98.12%; nickel recovery rate: 98.28%; molybdenum recovery rate: 98.36%; phosphorus recovery rate: 98.04%; vanadium recovery rate: 98.48%.

Acid leaching makes the molybdenum, phosphorus and vanadium in the alloy material not completely leached, and the content of molybdenum, phosphorus and vanadium in the recovered iron slag is more than 1%, and the filtrate after pressure acid leaching contains cobalt, nickel, molybdenum and vanadium metals. After the recovery of molybdenum vanadium and other metals by vanadium-absorbing resin and molybdenum-absorbing resin, it is necessary to recover cobalt and nickel through a further extraction process. The operation process is relatively complicated, and metal elements such as molybdenum and vanadium will also be dispersed in cobalt and nickel, resulting in The purity of cobalt and nickel metal is not high, and at the same time, the recovery rate of molybdenum and vanadium is also reduced.

Further, in order to achieve resource-limited treatment, the iron slag needs to be further leached with oxygen and alkali under pressure, and molybdenum, phosphorus, and vanadium are leached out to obtain high-purity iron slag and a molybdenum, phosphorus, and vanadium-containing leaching solution, and the iron slag can be effectively treated. .

In Example 1-Example 3, in the process of pressurized oxidizing alkali leaching, the comprehensive control of the pressure of the reaction kettle, the reaction temperature and the reaction time led to the adjustment of the degree of separation of each metal in the alloy material to the highest, Comparative Example 1 Compared with Example 1, the pressure and reaction temperature of the reactor are reduced, and the reaction time is prolonged. According to the results, it can be found that the leaching effect is not ideal. Compared with Example 1, Comparative Example 2 changed the parameters of the pressurized oxidative acid leaching process of iron, cobalt, and nickel filter residues, while Comparative Example 3 adopted the existing pressurized oxidative acid leaching method to separate the alloy material, and passed The data shows that the leaching rate of each metal in the alloy material is not high, and at the same time, the purity of each metal of iron, cobalt, and nickel is not high, which contains some impurities such as molybdenum and vanadium.

The experimental parameters of the above-mentioned Examples 1-3 and Comparative Example 1 are arranged in Table 1 as follows:

Table 1 is the data summary table of embodiment 1-5 and comparative example 1-3

*Acid consumption/alkali consumption is the total acid consumption/alkali consumption per kg of alloy material.

In the process method of the present invention, by adopting the method of pressurized oxidative alkali leaching and controlling the range of each parameter, the slurried alloy powder can react with alkali to generate corresponding molybdate, phosphoric acid and phosphoric acid under the coordination of various experimental parameters. Salt and vanadate are separated from iron slag, and the content of molybdenum, phosphorus and vanadium in iron slag is extremely low. The separation of molybdenum, phosphorus and vanadium by physical or chemical means has a remarkable effect. , Molybdenum, phosphorus and vanadium have achieved nearly complete leaching, and the purity of each metal is good. It has realized the full recovery and utilization of various metal resources, which is in line with the concept of sustainable development.

The above descriptions are only preferred embodiments of the present invention and are not intended to limit the present invention. Any modifications, equivalent replacements and improvements made within the spirit and principles of the present invention shall be included in the protection of the present invention. within the range.

Claims (10)

1. The leaching separation method of the alloy material is characterized by comprising the following steps:

s1, raw material treatment: preparing alloy waste into alloy powder; the alloy waste contains iron, cobalt, nickel, molybdenum, phosphorus and vanadium elements;

s2, batching and slurrying reaction: mixing alloy powder and alkaline solution according to a solid-liquid mass ratio of 1: 7-15, preparing a feed liquid, stirring at 70-90 ℃, and performing slurrying reaction for 1-5 h; the mass ratio of liquid alkali or flake alkali to alloy powder in the alkaline solution is 1.2-3.5: 1;

s3, pressure oxidation alkaline leaching: pumping the material liquid after the S2 reaction into a reaction kettle, introducing oxygen to increase the pressure to 1.8-2.8 MPa, and reacting for 2-8h at the temperature of 160-; after the reaction is finished, adjusting the pH value of the slurry to 7-9, and filtering to obtain filter residue and filtrate, wherein the filter residue contains iron, cobalt and nickel elements, and the filtrate contains molybdate ions, phosphate ions and vanadate ions;

s4, multi-metal separation: and (4) separating iron, cobalt and nickel elements from the filter residue obtained in the step S3, and separating molybdenum, phosphorus and vanadium elements from the filtrate obtained in the step S3.

2. The leaching separation method for the alloy material according to claim 1, wherein the separation process of molybdenum, phosphorus and vanadium elements in the S4 filtrate comprises the following steps:

s411, dephosphorization: crystallizing, washing and filtering the mixed filtrate obtained in the step S3 to obtain sodium phosphate crystals and a second filtrate, wherein the crystallization temperature is 1-5 ℃;

s412a, molybdenum separation: adjusting the pH value of the second filtrate obtained in the step S411 to 7.0-9.0, and adsorbing the second filtrate through molybdenum absorption resin at the temperature of 25-60 ℃ to obtain a third filtrate after molybdenum ions are removed; resolving the molybdenum adsorption resin to obtain a molybdate solution;

s413a, separation of vanadium: and adsorbing the third filtrate after molybdenum ion removal through vanadium-adsorbing resin, and analyzing the vanadium-adsorbing resin to obtain a vanadate solution.

3. The leaching separation method of an alloy material according to claim 2, wherein the resolving solution used in the resolving process in step S412a and step S413a is a sodium hydroxide solution with a mass fraction of 10-20%.

4. The leaching separation method of an alloy material according to claim 2, wherein the model of the molybdenum-adsorbing resin is any one of ZGD314, D352 and PDM.

5. The leaching and separating method of an alloy material as claimed in claim 2, wherein the model of the vanadium-absorbing resin is ZGD231 or LS-32.

6. The leaching separation method of alloy materials according to claim 2, wherein the second filtrate obtained in S411 is separated in the following way:

s412b, preparation of industrial ammonium metavanadate: adjusting the pH of the second filtrate to 7.0-9.5 by using an aqueous solution of ammonium sulfate with the mass fraction of 25% -35%, stirring at 20-75 ℃, and filtering to obtain industrial-grade ammonium metavanadate and fourth filtrate containing sodium molybdate;

s413b, and preparation of molybdenum trioxide: and (3) regulating the pH of the fourth filtrate containing sodium molybdate obtained in the step (S412) to 0.5-4.5 by using 60-70% by mass of perchloric acid, stirring at 30-80 ℃, filtering and drying to obtain molybdenum trioxide.

7. The leaching and separating method of the alloy material according to claim 1, wherein the step of separating iron, cobalt and nickel elements in the filter residue at S4 comprises the following steps:

s421, pressure oxidation acid leaching: and mixing the filter residue obtained in the step S3 with an acidic solution according to a solid-liquid mass ratio of 1: 6-15, preparing a feed liquid, stirring at 70-90 ℃, performing slurrying reaction for 1-5h, pumping the slurry into a reaction kettle, pressurizing to 1-2.8Mpa, reacting for 2-8h, and filtering to obtain iron slag and a fourth filtrate, wherein the fourth filtrate contains nickel ions and cobalt ions; the acid solution is sulfuric acid;

s422, extracting and separating nickel and cobalt elements, and sequentially carrying out chemical impurity removal and extraction on the fourth filtrate obtained in the S421 to obtain a cobalt sulfate solution and a nickel sulfate solution; the extractant is P507;

the mass ratio of the sulfuric acid in the acid solution to the filter residue is 1.5-3.5: 1.

8. the leaching separation method of the alloy material according to claim 6, wherein in S421, the pH of the solution after the reaction is adjusted to 2.0-4.0, and the solution is filtered to obtain the iron slag and a fourth filtrate.

9. The leaching separation method for the alloy material according to claim 1, wherein the filter residue obtained from the step S3 is subjected to reverse washing at least twice until the pH of the filtrate in the washing process is 7.0-8.0.

10. The method as claimed in claim 1, wherein the grain size of the alloy powder is 100-200 mesh.

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