Method for enriching precious metal from carbonylation alloy
Through the steps of electro-controlled chlorination and sulfation and roasting, the problems of decreasing precious metal content and increasing impurities in the carbonyl alloy are solved, and the effect of efficiently removing impurities and improving the content and purity of precious metals is achieved.
Patent Information
- Application Number
- CN202411504328.7
- Authority / Receiving Office
- CN · China
- Patent Type
- Applications(China)
- Current Assignee / Owner
- Filing Date
- 2024-10-25
- Publication Date
- 2025-02-25
AI Technical Summary
The content of precious metals in carbonyl alloys decreases, and the content of selenium, cobalt and sulfur increases, affecting the purification process of platinum, palladium and gold, resulting in difficulty in production operation and unstable product quality.
The impurities of copper sulfide and nickel sulfide in the carbonyl alloy are redoxified and decomposed by electrostatically controlled chlorination method, and the noble metal enters the slag and the base metal enters the liquid for recycling. Then the sulfation and calcination and water immersion steps are carried out to remove elemental sulfur and selenium, and the content and purity of precious metals are improved.
Effectively removes selenium, cobalt and sulfur impurities, improves the content and purity of precious metals in platinum concentrate, simplifies the process flow, reduces costs, and improves the direct yield and purification rate of precious metals.
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Figure CN119506590A_ABST
Abstract
Description
Technical Field
[0001] The present invention relates to the technical field of a process for enriching precious metals such as platinum, palladium, gold, rhodium and iridium by carbonylation alloy, and in particular to a method for enriching precious metals by carbonylation alloy. Background Art
[0002] With the adjustment of the industry, the raw materials for precious metal smelting have been transformed from secondary alloys to carbonyl alloys. Carbonyl alloys are water-quenched alloys produced by high-matte grinding and flotation. Under carbonylation conditions, nickel and iron generate gaseous carbonyls that are separated, while the precious metals in the alloy do not undergo carbonylation reaction and are almost all enriched in the carbonylation tailings. Compared with secondary alloys, the composition has changed significantly. The content of platinum, palladium and gold has decreased from the original 2000g / t to 700g / t, but the content of selenium, cobalt and sulfur has increased significantly. According to the process of processing secondary alloys to enrich carbonyl alloys, the selenium content in the produced platinum concentrate increased from the original 0.1% to 4-6%, the sulfur increased to 28%, and the cobalt content was as high as 16%. Because of the influence of sulfur, selenium and cobalt on the purification of platinum, palladium and gold precious metals, especially in the extraction process of platinum, palladium and gold, it seriously affects the production operation and product quality, and the direct recovery rate and grade rate of platinum, palladium and gold cannot be guaranteed. Therefore, a new method for enriching platinum concentrate with low selenium, sulfur and cobalt impurities and high precious metal content is particularly important.
[0003] Therefore, a method for enriching precious metals by carbonylation alloy is proposed. Summary of the invention
[0004] In view of the above technical problems, the present invention provides a method for enriching precious metals by carbonylating alloys.
[0005] In order to achieve the above object, the technical solution of the present invention is as follows: A method for enriching precious metals by carbonylation alloys comprises the following steps: S1. Carbonyl alloy with a particle size of 1mm-3mm and 2.0-2.5mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-to-solid ratio of 4-5:1. The agitators of the eight cells of the two mechanical kettles in series are turned on according to the position where the feed liquid arrives. At the same time, chlorine gas is introduced into a cell of the first mechanical kettle to increase the potential. In order to maintain the controlled electrochlorination time and chlorination effect, the potential of the two cells should be between 300 and 360mv. The temperature needs to adjust the steam flow rate to control the feed liquid temperature between 90℃ and 120℃. At the same time, the feed amount is adjusted to ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are between 390 and 420mv. The controlled electrochlorination time is maintained at 10-12h. The feed liquid after controlled electrochlorination is stored in a slurry tank and then filtered and separated to obtain filter residue and filtrate. The filtrate is returned to the copper system to recover copper. The filter residue is washed by slurry and then separated by filtration. The washing water is used for the hydrochloric acid preparation of controlled electrochlorination, and the slag enters the sulfuric acid roasting; S2, sulphate roasting: the slag after electrochlorination washing is added to concentrated sulphuric acid at a ratio of 2:1 into the rotary kiln. The rotary kiln can be divided into three temperature sections: the kiln head is the preheating and drying section with a temperature of 250-400°C; the kiln is the acidification roasting section with a temperature of 500-600°C; the kiln tail is the selenium steaming section with a temperature of 600-650°C. The material roasting time is 8-10 hours, and the negative pressure in the kiln is controlled at -147--200Pa; S3, water leaching: the material after sulfuric acid roasting is mixed with water at a liquid-to-solid ratio of 4-5:1, added into an enameled water leaching kettle, stirred and slurried, and steam is introduced for heating. The water leaching temperature is 80℃~93℃. After 0.5h of water leaching, it is filtered and separated. The filtrate is returned to the cobalt system for cobalt recovery, and the slag is used as platinum concentrate to enter the next purification process to separate platinum, palladium and gold. The content of precious metals in the platinum concentrate reaches 8-12%, the sulfur content is 8-15%, the nickel and copper content is 1-5%, the cobalt content is 1-3%, and the selenium content is 0.5-1%. This material enters the purification process for refining platinum, palladium and gold.
[0006] Wherein, in the step S1, the material is pushed into the mechanical kettle by a screw feeder, the frequency is 3-10 Hz, the feed amount is 0.3-0.5 t / h, and the amount of acid added is 0.75-1.1 L / h.
[0007] Wherein, in the step S1, the electro-chlorinated slag is slurried and washed three times and then separated by filter pressing.
[0008] Among them, in step S2, the gases produced by the sulfation roasting are selenium dioxide and sulfur dioxide, which pass through a four-stage water absorption tower to produce an oxidation-reduction reaction in water to generate more than 95% crude selenium, which is then filtered and separated to obtain a crude selenium product.
[0009] The beneficial effects of the present invention are: 1. The present invention utilizes controlled electrochlorination to oxidize and reduce the copper sulfide and nickel sulfide impurities in the carbonyl alloy at a certain potential. Sulfur enters the slag along with the precious metals, and base metals such as nickel, copper, and iron enter the liquid to recover the valuable metals. After the slag is washed and neutralized, the precious metals and elemental sulfur enter the rotary kiln and are added with concentrated sulfuric acid for sulfation roasting. After the elemental sulfur and selenium are removed, water is added to slurry and heated for water immersion. After a period of time, the liquid is separated by filter pressing. The liquid is used to recover valuable metals such as cobalt, and the slag is a platinum concentrate with qualified indicators, which can be sent to the purification process for the refining of platinum and palladium.
[0010] 2. The whole process of the present invention is simple to operate, with small amount of reagents, low cost and short production cycle. 90% of selenium can be removed while removing sulfur. Compared with the original process, a large amount of reagents used in sodium hyposulfurization and secondary chlorination are avoided. The operation is simple, the process flow is short, the cost is low, and the direct recovery rate of precious metals is high. It can provide necessary guarantee for the separation and refining of platinum, palladium and gold in the next purification process, and a certain amount of crude selenium by-product will be produced at the same time. BRIEF DESCRIPTION OF THE DRAWINGS
[0011] The present invention is further described below in conjunction with the accompanying drawings and embodiments: Figure 1 It is a process flow diagram of the present invention. DETAILED DESCRIPTION
[0012] In order to make the purpose, technical scheme and advantages of the present invention clearer, the present invention is further described in detail below in conjunction with specific embodiments and with reference to the accompanying drawings. It should be understood that these descriptions are only exemplary and are not intended to limit the scope of the present invention. In addition, in the following description, the description of well-known structures and technologies is omitted to avoid unnecessary confusion of the concept of the present invention.
[0013] The working principle of the present invention is: Hydrochloric acid electrochlorination: The controlled potential selective chlorination of base metals is to utilize the difference in chlorination potential of base metals such as copper, nickel, and iron. By controlling the consumption of chlorine in the hydrochloric acid medium, the solution is kept at a certain chlorination potential value. At this potential, the base metal undergoes chlorination reaction and enters the solution, while the precious metal does not react and remains in the slag, thereby achieving the purpose of separating the precious and base metals. The reaction equation is: Leaching: Cu+Cl 2 = CuCl 2 Ni+Cl 2 = NiCl 2 Ni+2CuCl 2 = NiCl 2 +2CuCl Chlorination: 2CuCl+Cl 2 = 2CuCl 2 CuS+Cl 2 = CuCl 2 +S Ni 2 S 3 +6CuCl 2 = 3NiCl 2 +6CuCl 2 +2S↓ Ni 3 S 2 +6Cl 2 = 3NiCl 2 +2S↓ NiS+2CuCl 2 = NiCl 2 +2CuCl+S↓ NiS+Cl 2 = NiCl2 +S↓ Ni 3 S 2 +CuCl 2 = NiCl 2 +Cu+2NiS↓ Ni 3 S 2 +2CuCl 2 = 2NiCl 2 +Cu 2 S+NiS↓ Ni 3 S 2 +3CuCl 2 = 3NiCl 2 +3CuS↓ Fe+Cl 2 = FeCl 2 FeS+Cl 2 = FeCl 2 +S↓ Among all metals, platinum group metals have a relatively positive oxidation-reduction potential value. There is a large difference between the standard oxidation-reduction potential of base metals. Therefore, by controlling the oxidation-reduction potential value of the system, base metals can be selectively oxidized and leached to enrich platinum group metals in the residue. The standard oxidation-reduction potentials of gold, platinum group metals and some base metals are shown in Table 1.
[0014] When chlorinated in a chloride solution, the oxidation-reduction potential of platinum group metals decreases due to their strong tendency to generate stable chloride complex ions, thereby reducing the standard potential difference between them and base metals. Therefore, the accuracy of the oxidation potential control in the chlorination system is required to be more stringent. The oxidation potential of platinum group metals in the chlorination system is shown in Table 2.
[0015] The controlled electric chlorination operation can be intermittent or continuous. When the leached material in the system consumes chlorine and hypochlorous acid, the potential decreases again. Therefore, by controlling the amount of chlorine introduced, the potential can be kept within a predetermined range, achieving the purpose of selectively leaching base metals. In continuous chlorination, the potential value of the system can be maintained by controlling the material addition rate and the flow rate of chlorine introduced.
[0016] Sulfate roasting: When selenium is oxidized and burned at a certain temperature, selenium reacts with oxygen at the oxidation temperature to generate volatile selenium dioxide gas, which enters the absorption tank and dissolves in water to form selenious acid solution. Sulfur reacts with oxygen under oxidation conditions to generate sulfur dioxide, which enters the absorption tower for reduction to obtain 90% crude selenium. Most of the high-boiling-point platinum, palladium, gold, nickel, and copper in the raw material crude selenium in the platinum concentrate do not volatilize at this temperature, and remain in the oxidation slag to become the platinum concentrate after selenium removal. Low-boiling-point impurities (S, As, Hg, etc.) volatilize into the absorption tank together with selenium dioxide, participate in the reaction to generate crude selenium, and the unreacted sulfur dioxide in the oxidation furnace absorption tank is discharged with the tail gas because it is insoluble in water, thereby achieving the purpose of removing selenium and reducing sulfur.
[0017] The reaction equation of sulfuric acid roasting is: Se+O 2 =SeO 2 ↑ S+O 2 =SO 2 ↑ SeO 2 +H 2 O=H 2 SeO 3 H 2 SeO 3 +2SO 2 +H 2 O=Se↓+H 2 SO 4 . CoO+H 2 SO 4 ==CoSO 4 +H 2 O CuS + 2H 2 SO 4 (concentrated) → CuSO 4 + SO 2 ↑ + S↓ + 2H 2 O; Water leaching: Dissolve all sulfates formed by base metals and sulfuric acid in water, especially cobalt sulfate.
[0018] Example 1 like Figure 1As shown: Step 1: Treat 6 tons of carbonyl alloy, carbonyl alloy with a particle size of 1mm~3mm and 2.5mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-solid ratio of 4.5:1, and the stirring of the two mechanical kettles in series is turned on in turn. Chlorine is introduced into a cell of the first mechanical kettle to increase the potential. The potential of the second cell is between 320 and 360mv, and the steam flow is adjusted. The temperature is controlled between 90℃ and 110℃ to ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are all between 390 and 420mv. The electrochlorination time is controlled to be 10h and then filtered and separated. The liquid contains nickel and copper at 120~180g / L. Return to the nickel or copper system to recover base metals, the slag is washed by slurry three times and then separated by filter pressing. The washing water is used to prepare hydrochloric acid for controlling electrochlorination. The slag rate of this process is 29.5% of the carbonyl alloy, and the slag water content is about 36%; the second step: the slag is added to the rotary kiln at a ratio of 2:1 with concentrated sulfuric acid, the temperature at the kiln head is 370℃; the temperature in the kiln is 570℃; the temperature at the kiln tail is 640℃, the material roasting time is 8h, and the negative pressure in the kiln is controlled at -160~-180Pa; the slag rate of this process is 21.2%; the third step: the material after sulfuric acid roasting and water are added to the enamel kettle at a liquid-solid ratio of 5:1 for water immersion, the water immersion temperature is 93℃, and the water immersion time is 0.5h before separation by filter pressing. The filtrate is returned to the cobalt system to recover cobalt, and a total of 127.8 kg of concentrate is produced. The total slag rate is 2.13%, the content of precious metals in the filter residue can reach 11.32%, and the direct recovery rates of platinum, palladium and gold are 99.99%, 99.97% and 94.9% respectively. The sulfur content is 13.56%, the nickel and copper content is 4.86%, the cobalt content is 1.23%, and the selenium content is 0.7%. This material enters the purification process to refine platinum, palladium and gold. The original process can produce 99.99% of precious metals, and the direct recovery rate can reach 98%.
[0019] Example 2 like Figure 1As shown: Step 1: Treat 7 tons of carbonyl alloy, carbonyl alloy with a particle size of 1mm~3mm and 2.3mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-solid ratio of 4.8:1, and the stirring of the two mechanical kettles in series is turned on in turn. Chlorine is introduced into a cell of the first mechanical kettle to increase the potential. The potential of the second cell is between 340 and 360mv, and the steam flow is adjusted. The temperature is controlled between 90℃ and 110℃. Ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are all between 390 and 418mv. Control the electrochlorination time to 10h and then filter and separate. The liquid contains nickel and copper between 125 and 186g / L The slag is washed three times by slurry and then separated by filter press. The washing water is used to prepare hydrochloric acid for electro-chlorination. The slag rate in this process is 30% of the carbonyl alloy and the slag water content is about 35%. The second step: the slag is added to the rotary kiln at a ratio of 2:1 with concentrated sulfuric acid. The temperature at the kiln head is 380°C; the temperature in the kiln is 580°C; the temperature at the kiln tail is 650°C, the material roasting time is 8.5h, and the negative pressure in the kiln is controlled at -160~-180Pa; the slag rate in this process is 20%; the third step: the material roasted with sulfuric acid and water are added to the enamel kettle at a liquid-solid ratio of 5:1 for water immersion. The water immersion temperature is 90°C and the water immersion time is 0.5h before separation by filter press. The filtrate is returned to the cobalt system to recover cobalt, and a total of 142.1 kg of concentrate is produced. The total slag rate is 2.03%, the content of precious metals in the filter residue can reach 11.64%, and the direct recovery rates of platinum, palladium and gold are 99.99%, 99.98% and 94.9% respectively. The sulfur content is 12.78%, the nickel and copper content is 5.16%, the cobalt content is 1.65%, and the selenium content is 0.43%. This material enters the purification process to refine platinum, palladium and gold. The original process can produce 99.99% of precious metals, and the direct recovery rate can reach 98.12%.
[0020] Example 3 like Figure 1As shown: Step 1: Treat 8 tons of carbonyl alloy, add carbonyl alloy with a particle size of 1mm~3mm and 2.4mol / L hydrochloric acid solution into the mechanical kettle at a liquid-solid ratio of 4.8:1, start the stirring of the two mechanical kettles in series in turn, introduce chlorine into one cell of the first mechanical kettle to increase the potential, the potential of the second cell is between 350 and 360mv, adjust the steam flow, control the temperature between 90℃ and 110℃, ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are between 390 and 420mv, control the electrochlorination time to 10h, and then filter and separate, the liquid contains nickel and copper between 110 and 172g / L, and return The nickel and copper systems recycle base metals. The slag is washed by slurry three times and then separated by filter pressing. The washing water is used to prepare hydrochloric acid for controlled electrochlorination. The slag rate of this process is 30.28% of the carbonyl alloy, and the slag water content is about 36.4%; the second step: the slag is added to the rotary kiln at a ratio of 2:1 with concentrated sulfuric acid. The temperature at the kiln head is 380°C; the temperature in the kiln is 590°C; the temperature at the kiln tail is 650°C, the material roasting time is 8.2h, and the negative pressure in the kiln is controlled at -170~-180Pa; the slag rate of this process is 20.12%; the third step: the material after sulfuric acid roasting and water are added to the enamel kettle at a liquid-solid ratio of 5:1 for water immersion. The water immersion temperature is 93°C, and the water immersion time is 0.5h before separation by filter pressing. The filtrate is returned to the cobalt system to recover cobalt, and a total of 164.8 kg of concentrate is produced. The total slag rate is 2.06%, the content of precious metals in the filter residue can reach 11.78%, and the direct recovery rates of platinum, palladium and gold are 99.98%, 99.97% and 94.8% respectively. The sulfur content is 13.02%, the nickel and copper content is 4.98%, the cobalt content is 1.23%, and the selenium content is 0.78%. This material enters the purification process to refine platinum, palladium and gold. The original process can produce 99.99% of precious metals, and the direct recovery rate can reach 98.11%.
[0021] Example 4 like Figure 1 As shown, a method for enriching precious metals by carbonylation alloy comprises the following steps: S1. Carbonyl alloy with a particle size of 1mm and 2.0mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-solid ratio of 4:1. The agitators of the eight cells of the two mechanical kettles in series are turned on according to the position where the feed liquid arrives. At the same time, chlorine gas is introduced into a cell of the first mechanical kettle to increase the potential. In order to maintain the controlled electrochlorination time and chlorination effect, the potential of the two cells should be between 300mv. The temperature needs to adjust the steam flow rate to control the feed liquid temperature between 90℃. At the same time, the feed amount is adjusted to ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are between 390mv. The controlled electrochlorination time is maintained at 12h. The feed liquid after controlled electrochlorination is stored in a slurry tank and then filtered and separated to obtain filter residue and filtrate. The filtrate is returned to the copper system to recover copper. The filter residue is washed by slurry and then separated by filtration. The washing water is used for the preparation of hydrochloric acid for controlled electrochlorination, and the slag enters the sulfuric acid roasting; S2, sulphate roasting: the slag after electrochlorination washing is added to concentrated sulphuric acid at a ratio of 2:1 into a rotary kiln. The rotary kiln can be divided into three temperature sections: the kiln head is a preheating and drying section with a temperature of 250°C; the kiln is an acidification roasting section with a temperature of 500°C; the kiln tail is a selenium steaming section with a temperature of 600°C. The material roasting time is 10 hours, and the negative pressure in the kiln is controlled at -147Pa; S3, water leaching: the material after sulfuric acid roasting is mixed with water at a liquid-to-solid ratio of 4:1, and added into an enameled water leaching kettle. The mixture is stirred and slurried, and steam is introduced for heating. The water leaching temperature is 80°C. After 0.5h of leaching, it is filtered and separated. The filtrate is returned to the cobalt system for cobalt recovery, and the slag is used as platinum concentrate to enter the next purification process to separate platinum, palladium and gold. The content of precious metals in the platinum concentrate reaches 8-12%, the content of sulfur is 8-15%, the content of nickel and copper is 1-5%, the content of cobalt is 1-3%, and the content of selenium is 0.5-1%. This material enters the purification process for refining platinum, palladium and gold.
[0022] Wherein, in step S1, the material is pushed into the mechanical kettle by a screw feeder, the frequency is 3 Hz, the feed rate is 0.3 t / h, and the amount of acid added is 0.75 L / h.
[0023] Wherein, in the step S1, the electro-chlorinated slag is slurried and washed three times and then separated by filter pressing.
[0024] Among them, in step S2, the gases produced by the sulfation roasting are selenium dioxide and sulfur dioxide, which pass through a four-stage water absorption tower to produce an oxidation-reduction reaction in water to generate more than 95% crude selenium, which is then filtered and separated to obtain a crude selenium product.
[0025] Example 5 like Figure 1 As shown, a method for enriching precious metals by carbonylation alloy comprises the following steps: S1. Carbonyl alloy with a particle size of 1mm-3mm and 2.0mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-to-solid ratio of 5:1. The agitators of the eight cells of the two mechanical kettles in series are turned on according to the position where the feed liquid arrives. At the same time, chlorine gas is introduced into a cell of the first mechanical kettle to increase the potential. In order to maintain the controlled electrochlorination time and chlorination effect, the potential of the two cells should be between 360mv. The temperature needs to adjust the steam flow rate to control the feed liquid temperature between 120℃. At the same time, the feed amount is adjusted to ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are between 420mv. The controlled electrochlorination time is maintained at 10h. The feed liquid after controlled electrochlorination is stored in a slurry tank and then filtered to obtain filter residue and filtrate. The filtrate is returned to the copper system to recover copper. The filter residue is washed by slurry and then filtered. The washing water is used for the hydrochloric acid preparation of controlled electrochlorination, and the slag enters the sulfuric acid roasting; S2, sulphate roasting: the slag after electrochlorination washing is added to concentrated sulphuric acid at a ratio of 2:1 into the rotary kiln. The rotary kiln can be divided into three temperature sections: the kiln head is the preheating and drying section with a temperature of 400°C; the kiln is the acidification roasting section with a temperature of 600°C; the kiln tail is the selenium steaming section with a temperature of 650°C. The material roasting time is 8 hours, and the negative pressure in the kiln is controlled at -200Pa; S3, water leaching: the material after sulfuric acid roasting is mixed with water at a liquid-to-solid ratio of 5:1, and added into an enameled water leaching kettle. The mixture is stirred and slurried, and steam is introduced for heating. The water leaching temperature is 93°C. After 0.5h of leaching, it is filtered and separated. The filtrate is returned to the cobalt system for cobalt recovery, and the slag is used as platinum concentrate to enter the next purification process to separate platinum, palladium and gold. The content of precious metals in the platinum concentrate reaches 8-12%, the content of sulfur is 8-15%, the content of nickel and copper is 1-5%, the content of cobalt is 1-3%, and the content of selenium is 0.5-1%. This material enters the purification process for refining platinum, palladium and gold.
[0026] Wherein, in step S1, the material is pushed into the mechanical kettle by a screw feeder, the frequency is 10 Hz, the feed rate is 0.5 t / h, and the amount of acid added is 1.1 L / h.
[0027] Wherein, in the step S1, the electro-chlorinated slag is slurried and washed three times and then separated by filter pressing.
[0028] Among them, in step S2, the gases produced by the sulfation roasting are selenium dioxide and sulfur dioxide, which pass through a four-stage water absorption tower to produce an oxidation-reduction reaction in water to generate more than 95% crude selenium, which is then filtered and separated to obtain a crude selenium product.
[0029] It should be understood that the above specific embodiments of the present invention are only used to illustrate or explain the principles of the present invention, and do not constitute a limitation of the present invention. Therefore, any modifications, equivalent substitutions, improvements, etc. made without departing from the spirit and scope of the present invention should be included in the protection scope of the present invention. In addition, the appended claims of the present invention are intended to cover all changes and modifications that fall within the scope and boundaries of the appended claims, or the equivalent forms of such scope and boundaries.
Claims
1. A method for enriching precious metals by carbonylation alloy, characterized in that: The following steps are involved: S1. Carbonyl alloy with a particle size of 1mm-3mm and 2.0-2.5mol / L hydrochloric acid solution are added to the mechanical kettle at a liquid-to-solid ratio of 4-5:
1. The agitators of the eight cells of the two mechanical kettles in series are turned on according to the position where the feed liquid arrives. At the same time, chlorine gas is introduced into a cell of the first mechanical kettle to increase the potential. In order to maintain the controlled electrochlorination time and chlorination effect, the potential of the two cells should be between 300 and 360mv. The temperature needs to adjust the steam flow rate to control the feed liquid temperature between 90℃ and 120℃. At the same time, the feed amount is adjusted to ensure that the potential of the fourth cell of the 1# mechanical kettle and the four cells of the 2# kettle connected in series are between 390 and 420mv. The controlled electrochlorination time is maintained at 10-12h. The feed liquid after controlled electrochlorination is stored in a slurry tank and then filtered and separated to obtain filter residue and filtrate. The filtrate is returned to the copper system to recover copper. The filter residue is washed by slurry and then separated by filtration. The washing water is used for the hydrochloric acid preparation of controlled electrochlorination, and the slag enters the sulfuric acid roasting; S2, sulphate roasting: the slag after electrochlorination washing is added to concentrated sulphuric acid at a ratio of 2:1 into the rotary kiln. The rotary kiln can be divided into three temperature sections: the kiln head is the preheating and drying section with a temperature of 250-400°C; the kiln is the acidification roasting section with a temperature of 500-600°C; the kiln tail is the selenium steaming section with a temperature of 600-650°C. The material roasting time is 8-10 hours, and the negative pressure in the kiln is controlled at -147--200Pa; S3, water leaching: the material after sulfuric acid roasting is mixed with water at a liquid-to-solid ratio of 4-5:1, added into an enameled water leaching kettle, stirred and slurried, and steam is introduced for heating. The water leaching temperature is 80℃~93℃. After 0.5h of water leaching, it is filtered and separated. The filtrate is returned to the cobalt system for cobalt recovery, and the slag is used as platinum concentrate to enter the next purification process to separate platinum, palladium and gold. The content of precious metals in the platinum concentrate reaches 8-12%, the sulfur content is 8-15%, the nickel and copper content is 1-5%, the cobalt content is 1-3%, and the selenium content is 0.5-1%. This material enters the purification process for refining platinum, palladium and gold.
2. The method for enriching precious metals by carbonylation alloy according to claim 1, characterized in that: In the step S1, the material is pushed into the mechanical kettle by a screw feeder, the frequency is 3-10 Hz, the feed amount is 0.3-0.5 t / h, and the amount of acid added is 0.75-1.1 L / h.
3. The method for enriching precious metals by carbonylation alloy according to claim 1, characterized in that: In the step S1, the electro-chlorinated slag is slurried and washed three times and then separated by filter pressing.
4. The method for enriching precious metals by carbonylation alloy according to claim 1, characterized in that: In step S2, the gases generated by the sulfation roasting are selenium dioxide and sulfur dioxide, which pass through a four-stage water absorption tower to produce an oxidation-reduction reaction in water to generate more than 95% crude selenium, which is then filtered and separated to obtain a crude selenium product.