A method for enriching precious metals from carbonylation tailings

Abstract

The present invention relates to the technical field of metal separation and purification, and more particularly to a method for enriching precious metals from carbonylation tailings. The method comprises smelting and reduction of the carbonylation tailings, slow cooling and crystallization, grinding and classification, controlled electrochlorination, sodium sulfate desulfurization, and secondary controlled electrochlorination, ultimately producing a precious metal-enriched concentrate capable of separating and purifying the precious metals. The method is simple in process, has a high enrichment ratio and recovery rate for rare and precious metals, can separate and recover nickel and copper, and achieves a high comprehensive resource utilization rate.

Description

Technical Field

[0001] The invention relates to the technical field of metal separation and purification, and in particular to a method for enriching precious metals from carbonylation tailings. Background Art

[0002] During the smelting process of nickel sulfide-associated platinum group metal ores, the nickel ore raw material undergoes flash smelting to form low-grade nickel matte. The low-grade nickel matte is then blown to produce high-grade nickel matte. After slow cooling, crystallization, grinding, and sorting, the primary nickel and copper concentrates and primary alloys are produced. The primary alloy and hot filter residue are sulfurized in an alloy sulfurization furnace, slow cooling, crystallization, and further grinding and sorting to produce secondary nickel concentrate, secondary copper concentrate, and secondary alloys. During each slow cooling, crystallization, and subsequent grinding, approximately 70% of the platinum group metals enter the alloy phase. The final amount of platinum group metals entering the secondary alloy is less than 50% of the amount introduced into the raw material, resulting in a low direct yield of platinum group metals. Therefore, after years of research, a carbonylation production process was developed, in which the primary alloy is water-quenched to form a water-quenched alloy, which then enters the nickel carbonyl production system. Under the carbonylation reaction conditions, the water-quenched alloy forms a gaseous carbonyl from nickel and iron, which is separated and used to produce high-purity nickel and iron. The precious metals in the alloy do not undergo carbonylation and are almost entirely concentrated in the carbonylation tailings. The solid residue after carbonylation extraction of nickel and iron is called carbonylation tailings. The copper content in carbonylation tailings is 60~65%, the sulfur content is 16~18%, and the nickel content is about 8~15%. In addition, in the process of carbonylation synthesis of carbonyl nickel and iron, the original nickel-copper alloy lattice is broken, resulting in enhanced activity of precious metals. In the process of separating base metals by wet leaching process, the loss of precious metals is large. Therefore, new technologies need to be developed to adapt to new materials. Summary of the Invention

[0003] In view of the problems existing in the above-mentioned background technology, the present invention proposes a method for enriching precious metals from carbonyl alloys. The method has the advantages of simple process, short flow and high precious metal yield.

[0004] The present invention adopts the following technical solutions:

[0005] A method for enriching precious metals from carbonylation tailings comprises the following steps:

[0006] Step a: After uniformly mixing the carbonylation tailings, coke and sodium carbonate, the mixture is placed in a muffle furnace and heated to 1150-1250°C for melting. After the materials are completely melted, the mixture is allowed to stand for 1-2 hours for smelting reduction. The amount of coke added is 0.5%-1% of the amount of the carbonylation tailings, and the amount of sodium carbonate added is 1%-2% of the amount of the carbonylation tailings.

[0007] Step b: slowly cooling and crystallizing the melt obtained in step a in a muffle furnace for 72 hours to obtain a crude copper alloy;

[0008] Step c: grinding and screening the slowly cooled and crystallized crude copper alloy using a Raymond mill to obtain a crude copper alloy with a particle size of 180-200 mesh;

[0009] Step d: adding the crude copper alloy to 2.5-3.5 mol / L hydrochloric acid, controlling the liquid-solid ratio to be 5-6:1, heating to 80-90° C., and introducing chlorine gas for controlled electrochlorination leaching; after raising the pulp potential to 395-405 mV, turning off the chlorine gas, continuing to raise the temperature to 90-95° C., maintaining the temperature for 2-3 hours, and performing constant potential leaching; after the reaction is completed, adding gelatin and continuing the reaction for 30 minutes, filtering to obtain a leaching residue and a leachate, the leachate is sent to a copper salt plant to produce copper sulfate and separate and recover metallic nickel, and the leaching residue is sent to the next step for treatment;

[0010] Step e: adding water to the leaching residue obtained in step d to slurry, controlling the liquid-solid ratio to be 6-7:1, adding caustic soda flakes, adjusting the pH to 8-9, adding anhydrous sodium sulfite in an amount of 4.5-5 times the mass of the sulfur in the leaching residue, heating to a boiling state, and maintaining for 20-35 minutes, stopping heating, and maintaining the temperature at 90-95° C., adding copper powder in an amount of 0.2-0.3% the mass of the leaching residue, and continuing to react for 20-40 minutes. After the reaction is completed, filtering to obtain desulfurized residue and desulfurized liquid; the desulfurized residue is transferred to the next step for treatment, and the desulfurized liquid is sold to produce baking soda;

[0011] Step f: adding the desulfurization slag obtained in step e to 2-2.5 mol / L hydrochloric acid, controlling the liquid-solid ratio to 3-4:1, heating to 80-90° C., introducing chlorine gas for secondary controlled electrochlorination leaching; after increasing the pulp potential to 380-385 mv, maintaining the potential and continuing leaching for 6-12 hours, adding gelatin after the reaction is completed and continuing the reaction for 30 minutes, filtering to obtain a concentrate and a secondary leachate, handing the concentrate over to a precious metal purification process for separation and extraction of precious metals, and the secondary leachate can be reused for controlled electrochlorination.

[0012] The recovery principle of the present invention is as follows: a primary alloy is treated by the carbonylation method. Under high-temperature, high-pressure carbonylation conditions, Ni is released from the Cu-Ni alloy to produce Ni(CO)4. The release of nickel in the resulting carbonylation alloy alters the original Cu-Ni alloy crystal structure compared to before carbonylation, exposing the platinum group metals embedded in the Cu-Ni alloy and enhancing the activity of the platinum group metals. Therefore, during the wet leaching process, some of the platinum group precious metals enter the solution, resulting in precious metal loss. The present invention reconstructs the residual copper lattice by a smelting reduction method, re-encapsulating the precious metals within the copper lattice structure. The crystal structure then grows through a slow cooling crystallization process, thereby reducing the activity of the precious metals and minimizing precious metal dissolution losses during the chlorination leaching process.

[0013] Compared with the prior art, the present invention has the following beneficial effects:

[0014] The method of the present invention can efficiently leach and recover the valuable metals nickel and copper from carbonylation tailings, while simultaneously enriching precious metals. The resulting copper- and nickel-rich leachate can be used as a raw material for nickel and copper extraction. The leached residue is enriched with platinum-group metals and then transferred to a platinum-group metal production line for refining. After pyrometallurgical smelting, the crystal lattice of the carbonylation tailings is reconstructed, significantly reducing the activity of the precious metals and improving the yield of precious metals during the treatment process. The nickel and copper leaching rate in the carbonylation alloy can be increased to over 99%, while the slag rate is 2-3%. The precious metals are enriched by more than 30 times, facilitating subsequent refining and purification of the platinum-group metals. BRIEF DESCRIPTION OF THE DRAWINGS

[0015] Figure 1 The process flow chart of the present invention is DETAILED DESCRIPTION

[0016] The raw material carbonylation tailings used are shown in Table 1 below:

[0017] Table 1 Element content in carbonylation tailings (unit: precious metal g / t, other %)

[0018]

[0019] The carbonylation tailings were mixed evenly with 0.5% to 1% coke and 1% to 2% sodium carbonate by mass, placed in a muffle furnace, and heated to 1150-1250°C for melting. After the material was completely melted, it was allowed to stand for 1-2 hours for smelting reduction. The resulting melt was slowly cooled and crystallized in a muffle furnace for 72 hours to obtain a crude copper alloy. The crude copper alloy after slow cooling and crystallization was ground and sieved using a Raymond mill to obtain a crude copper alloy with a particle size of 180-200 mesh. The crude copper alloy yield was 93%, with no smelting slag produced. There was almost no loss of gold, platinum, palladium, nickel, and copper in this stage. The weight loss was mainly sulfur, accounting for about 15% of the total sulfur. The composition of the crude copper alloy is shown in Table 2 below:

[0020] Table 2 Element content in crude copper alloy (unit: precious metal g / t, other %)

[0021]

[0022] Example 1

[0023] The ground copper alloy was added to 2.5 mol / L hydrochloric acid, with a liquid-to-solid ratio of 5:1. After heating to 80°C, chlorine gas was introduced for controlled electrochlorination leaching. After the pulp potential was raised to 395 mV, the chlorine gas was turned off and the temperature was continued to rise to 90°C. The temperature was kept constant for 2 hours for constant potential leaching. After the reaction is completed, gelatin is added to continue the reaction for 30 minutes, and the mixture is filtered to obtain leaching residue; the leaching residue is pulped, the liquid-solid ratio is controlled to 6:1, caustic soda is added, the pH is adjusted to 8, and anhydrous sodium sulfite accounting for 4.5 times the mass of sulfur in the leaching residue is added, and the mixture is heated to boiling state and maintained for 20 minutes. The heating is stopped, the temperature is kept constant at 90°C, and copper powder accounting for 0.2% of the mass of the leaching residue is added to continue the reaction for 20 minutes. After the reaction is completed, the mixture is filtered to obtain desulfurized residue; the desulfurized residue is added to 2mol / L hydrochloric acid, the liquid-solid ratio is controlled to 3:1, and after heating to 80°C, chlorine is introduced for secondary controlled electrochlorination leaching. After the pulp potential is increased to 380mv, the potential is maintained and the leaching is continued for 6 hours. After the reaction is completed, gelatin is added to continue the reaction for 30 minutes, and the mixture is filtered to obtain precious metal concentrate. The specific results are shown in Table 3:

[0024] Table 3 Analysis of experimental results of Example 1 (solid:%, g / t, liquid: g / L)

[0025]

[0026] Example 2

[0027] The ground copper alloy was added to 3.5 mol / L hydrochloric acid, with a liquid-to-solid ratio of 6:1. After heating to 90°C, chlorine gas was introduced for controlled electrochlorination leaching. After the pulp potential was raised to 405 mV, the chlorine gas was turned off and the temperature was continued to rise to 95°C. The temperature was kept constant for 3 hours for constant potential leaching. After the reaction is completed, gelatin is added to continue the reaction for 30 minutes, and the mixture is filtered to obtain leaching residue; the leaching residue is pulped, the liquid-solid ratio is controlled to 7:1, caustic soda is added, the pH is adjusted to 9, and anhydrous sodium sulfite is added that is 5 times the sulfur content in the leaching residue, and the mixture is heated to boiling and maintained for 35 minutes. The heating is then stopped, the temperature is kept constant at 95°C, and copper powder that is 0.3% of the mass of the leaching residue is added to continue the reaction for 40 minutes. After the reaction is completed, the mixture is filtered to obtain desulfurized residue; the desulfurized residue is added to 2.5 mol / L hydrochloric acid, the liquid-solid ratio is controlled to 4:1, and after heating to 90°C, chlorine is introduced for secondary controlled electrochlorination leaching. After the pulp potential is increased to 385 mv, the potential is maintained and the leaching is continued for 12 hours. After the reaction is completed, gelatin is added to continue the reaction for 30 minutes, and the mixture is filtered to obtain precious metal concentrate. The specific results are shown in Table 4:

[0028] Table 4 Experimental results analysis of Example 2 (solid:%, g / t, liquid: g / L)

[0029]

Claims

1. A method for enriching precious metals from carbonylation tailings, characterized in that: The following steps are involved: Step a: After uniformly mixing the carbonylation tailings, coke and sodium carbonate, the mixture is placed in a muffle furnace and heated to 1150-1250°C for melting. After the materials are completely melted, the mixture is allowed to stand for 1-2 hours for smelting reduction. The amount of coke added is 0.5%-1% of the mass of the carbonylation tailings, and the amount of sodium carbonate added is 1%-2% of the mass of the carbonylation tailings. Step b: slowly cooling and crystallizing the melt obtained in step a in a muffle furnace for 72 hours to obtain a crude copper alloy; Step c: grinding and screening the slowly cooled and crystallized crude copper alloy using a Raymond mill to obtain a crude copper alloy with a particle size of 180-200 mesh; Step d: adding the crude copper alloy to 2.5-3.5 mol / L hydrochloric acid, controlling the liquid-solid ratio to be 5-6:1, heating to 80-90° C., and introducing chlorine gas for controlled electrochlorination leaching; after raising the pulp potential to 395-405 mV, turning off the chlorine gas, continuing to raise the temperature to 90-95° C., maintaining the temperature for 2-3 hours, and performing constant potential leaching; after the reaction is completed, adding gelatin and continuing the reaction for 30 minutes, filtering to obtain a leaching residue and a leachate, the leachate is sent to a copper salt plant to produce copper sulfate and separate and recover metallic nickel, and the leaching residue is sent to the next step for treatment; Step e: adding water to the leaching residue obtained in step d to slurry, controlling the liquid-solid ratio to be 6-7:1, adding caustic soda flakes, adjusting the pH to 8-9, adding anhydrous sodium sulfite in an amount of 4.5-5 times the mass of the sulfur in the leaching residue, heating to a boiling state, and maintaining for 20-35 minutes, stopping heating, and maintaining the temperature at 90-95° C., adding copper powder in an amount of 0.2-0.3% the mass of the leaching residue, and continuing to react for 20-40 minutes. After the reaction is completed, filtering to obtain desulfurized residue and desulfurized liquid; the desulfurized residue is transferred to the next step for treatment, and the desulfurized liquid is sold to produce baking soda; Step f: adding the desulfurization slag obtained in step e to 2-2.5 mol / L hydrochloric acid, controlling the liquid-solid ratio to 3-4:1, heating to 80-90° C., introducing chlorine gas for secondary controlled electrochlorination leaching; after increasing the pulp potential to 380-385 mv, maintaining the potential and continuing leaching for 6-12 hours, after the reaction is completed, adding gelatin and continuing the reaction for 30 minutes, filtering to obtain a concentrate and a secondary leachate, handing the concentrate over to a precious metal purification process for separation and extraction of precious metals, and recycling the secondary leachate to step d for controlled electrochlorination.

2. A method for enriching precious metals from carbonylation tailings according to claim 1, characterized in that: The carbonylation tailings mainly comprise Au: 150-260 g / t; Pd: 240-350 g / t; Pt: 440-550 g / t; Ni: 8-15%; Cu: 60-65%; and S: 15-20%.

3. The method for enriching precious metals from carbonylation tailings according to claim 1, wherein: The main components of the crude copper alloy are Au: 165-275 g / t; Pd: 280-390 g / t; Pt: 570-680 g / t; Ni: 9-17%; Cu: 67-73%; and S: 13-17%.

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