A method for recovering precious metals from precious metal alloys

Abstract

The present invention belongs to the technical field of metal recycling, and a method for recycling precious metals from precious metal alloys includes the following steps: Step 1: Mix and melt an alloy containing precious metals and a fragmentation agent metal, and perform fragmentation treatment to obtain brittle alloy particles; Step 2: Mix and react the brittle alloy particles with an acid solution to obtain a solution and an acid leaching residue; the reaction activity of the fragmentation agent metal in the brittle alloy particles with the acid is higher than the reaction activity of the precious metals in the brittle alloy particles with the acid; at least part of the fragmentation agent metal in the acid leaching residue exists in the form of a single substance; Step 3: Completely oxidize the fragmentation agent metal in the acid leaching residue to obtain a mixed material; Step 4: Perform refining treatment on the mixed material to extract a precious metal compound or a precious metal single substance. This method can effectively remove active fragmentation agent metals while reducing the generation of wastewater, and reduce the loss of precious metals, thereby improving the direct recovery rate of precious metals.

Description

Technical Field

[0001] The present invention relates to the technical field of metal recycling, and particularly relates to a method for recycling precious metals from precious metal alloys. Background Art

[0002] Cemented carbide is made of refractory metal hard compounds and binding metals through powder metallurgy. Cemented carbide has a series of excellent properties such as high hardness, good wear resistance, high strength, heat resistance, and corrosion resistance. Especially its high hardness and wear resistance remain basically unchanged even at a temperature of 500 °C and still have a high hardness at 1000 °C. However, with the increase in the hardness of cemented carbide, its fracture toughness shows a downward trend. In order to meet the higher requirements in the use of mining tools, high-speed wire rolling mills, stamping dies, and non-ferrous metal processing, and at the same time improve the hardness and fracture toughness of cemented carbide, metals such as ruthenium, tungsten, cobalt, chromium, titanium, nickel, molybdenum, tantalum, and iron are usually added to it.

[0003] Ruthenium-based alloy materials are one of the important materials in vertical magnetic storage systems. For example, ruthenium thin films, as intermediate layers, play roles such as increasing thermal stability, reducing lattice mismatch stress, and reducing noise; ruthenium-based alloy thin films, as coupling layers in antiferromagnetic coupling magnetic recording media, have the characteristics of high recording density and high performance; RuCo / ReCo thin films, as particle layers isolating Ru-containing antiferromagnetic layers, are beneficial to increasing the areal recording density.

[0004] Ruthenium-containing cemented carbide or ruthenium-based alloy materials are collectively referred to as ruthenium alloys. The scraps, defective products, and other waste residues and chips in the production and manufacturing process are important secondary resources and have extremely high recycling value. In particular, the precious metal ruthenium in them must be recycled.

[0005] Ruthenium is difficult to dissolve with acids, alkalis, and oxidants, and even aqua regia cannot dissolve it under conventional conditions. The efficient and rapid dissolution of ruthenium is a recognized problem in the field of precious metal metallurgy. How to effectively dissolve metallic ruthenium into the solution is a key link in the field of ruthenium metallurgy and an important measure to improve the recovery rate. Due to the addition of some elements such as Re, Ta, W, Mo, and Co, ruthenium alloys have good corrosion resistance and oxidation resistance, and their hardness is very high, up to 700 HV. Large pieces of ruthenium alloys are more stable in nature, and the direct leaching efficiency is very low; their hardness exceeds various crushing equipment on the market, and the wear of the device during the crushing process is very large, and it is impossible to efficiently crush the materials to the qualified particle size, resulting in the difficult and efficient recycling of the precious metal ruthenium in the alloy.

[0006] Chinese Patent 01131879.1 discloses a method for extracting osmium, iridium, and ruthenium. The materials are fragmented with zinc-aluminum alloy, quenched with water, acid-dissolved and washed with water, and then purified through subsequent processes to obtain sponge ruthenium, and the direct recovery rate of ruthenium is 81.29%.

[0007] Chinese Invention Patent No. 201310234311.X also discloses a process for extracting precious metals from ruthenium-iridium-osmium ore: after the ruthenium-iridium-osmium ore is fragmented, most of the Os is oxidized and volatilized by pyrometallurgy and absorbed with NaOH solution. The residue is melted with sodium peroxide and then leached with water. Ru is reduced with ethanol, and the leaching residue is sent for iridium extraction. It is recorded in its specification: (1) Fragmentation: Put Zn and Al into a graphite crucible according to Zn∶Al (mass ratio) = 3-4∶1-2, place it in a muffle furnace and heat it to 500-800 °C to melt. Put 1 part of ruthenium-iridium-osmium ore wrapped in aluminum foil into the above melt, keep it warm for 2 h, and take it out and stir 2 times to make it melt evenly, then pour it into a stainless steel container for water quenching. Take out the water-quenched alloy block and gradually dissolve it with acid. The acid consumption is about 2 times the theoretical amount, the temperature is 30-90 °C, and the time is 24 h. The filtered precious metal powder is washed with water until neutral for further treatment. (2) Pyrometallurgical oxidation and volatilization: In this process, the fragmented ruthenium-iridium-osmium ore powder is oxidized and volatilized of Os at high temperature. Appropriately control the flow rate of oxygen or air, and Os is oxidized to OsO4 gas and volatilized and recovered. The specific operation is as follows: First, connect the absorption system of Os. Load a certain amount of precious metal powder into a quartz boat, dry it at 100 °C in a quartz tube electric furnace for 2 h, and then gradually heat it up to 300-800 °C and introduce air for oxidation. At the same time, turn on the vacuum pump to suck the generated OsO4 gas into the sodium hydroxide solution. This process lasts for 24 h, and a large amount of white gas emerges until it is tested that there is no OsO4 gas emerging. Then cut off the power and cool it to room temperature, and take out the material for the next step of treatment. During the oxidation and volatilization process, Ru is oxidized to RuO2.

[0008] The technical solutions of the above two patents both form alloys with materials at high temperature using zinc, aluminum or zinc-aluminum alloy; then dissolve zinc or aluminum with hydrochloric acid or sulfuric acid to obtain precious metal powder, so that the recovered precious metal powder has high purity and recovery rate; however, due to the development of technology, there are further requirements for the recovery purity and recovery rate of precious metals in the current market, and the technical solutions of the above two patents now have the following defects:

[0009] (1) The consumption of fragmentation agents, namely aluminum and zinc metals, is relatively large, which is 2.3-6 times that of the material, and there is a risk of explosion during the water quenching of molten aluminum.

[0010] (2) The consumption of acid is large, which is 1.2-2 times the theoretical amount (i.e., the acid equivalent). The dissolution rate of acid is slow. It takes 10-20 h to completely leach zinc and aluminum with acid, with low efficiency and incomplete acid leaching of zinc and aluminum.

[0011] (3) The activity of the precious metals after fragmentation is very strong, and there is a large loss of precious metals during the acid dissolution process, which requires re-recovery and affects the direct recovery rate of precious metals.

[0012] (4) It is difficult to completely separate the solid from the liquid after acid dissolution. The solid residue after acid leaching is acidic and requires a large amount of water and time to wash the acidic solid residue to neutrality, generating a large amount of wastewater and increasing the cost of wastewater treatment.

[0013] Therefore, it is necessary to develop a method for recovering precious metals with relatively low cost, high safety, and high direct recovery rate of precious metals. Summary of the Invention

[0014] The purpose of the present invention is to provide a method for recovering precious metals from a precious metal alloy, which can effectively remove active fragmentation agent metals while reducing the generation of wastewater, reducing the loss of precious metals, and improving the direct recovery rate of precious metals.

[0015] Based on the above purpose, the present invention provides a method for recovering precious metals from a precious metal alloy, including the following steps:

[0016] Step 1: Mix and melt an alloy containing precious metals and a fragmentation agent metal, and perform fragmentation treatment to obtain brittle alloy particles;

[0017] Step 2: Mix and react the brittle alloy particles with an acid solution to obtain a solution and an acid leaching residue; the reaction activity of the fragmentation agent metal in the brittle alloy particles with the acid is higher than that of the precious metal in the brittle alloy particles with the acid; at least part of the fragmentation agent metal in the acid leaching residue exists in the form of a single substance;

[0018] Step 3: Completely oxidize the fragmentation agent metal in the acid leaching residue to obtain a mixed material;

[0019] Step 4: Perform refining treatment on the mixed material to extract and obtain a precious metal compound or a precious metal single substance.

[0020] Preferably, the precious metal is at least one of ruthenium, osmium, and iridium.

[0021] Preferably, in Step 1, the dosage of the fragmentation agent metal is at least 1 times the weight of the alloy containing precious metals.

[0022] More preferably, in Step 1, the dosage of the fragmentation agent metal is at least 1.1 - 1.4 times the weight of the alloy containing precious metals.

[0023] More preferably, in Step 1, the dosage of the fragmentation agent metal is at least 1.1 times, 1.2 times, 1.3 times, or 1.4 times the weight of the alloy containing precious metals.

[0024] Preferably, the fragmentation agent metal is aluminum or / and zinc.

[0025] Furthermore, the fragmentation agent metal is preferably aluminum and zinc, but aluminum-zinc mixtures are not excluded, and the aluminum-zinc mixtures can be mixed in any proportion.

[0026] Preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 1 wt%-25 wt% of the total amount of the fragmentation agent metal.

[0027] More preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 5 wt%-20 wt% of the total amount of the fragmentation agent metal.

[0028] More preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 7 wt%-18 wt% of the total amount of the fragmentation agent metal.

[0029] More preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 9 wt%-16 wt% of the total amount of the fragmentation agent metal.

[0030] More preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 11 wt%-15 wt% of the total amount of the fragmentation agent metal.

[0031] Preferably, in the step 2, the fragmentation agent metal in the form of simple substance in the acid leaching residue is equivalent to 5 wt%, 7 wt%, 9 wt%, 11 wt%, 13 wt%, 15 wt%, 17 wt%, 19 wt% of the total amount of the fragmentation agent metal.

[0032] Preferably, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:5-8; the acid in the acid solution is sulfuric acid or hydrochloric acid; the reaction time is 4-6 h, and the temperature is 20-90 °C.

[0033] Further, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:5.5-7.5.

[0034] Further, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:6-7.

[0035] Further, in the step 2, the reaction time is 3-5 h, and the temperature is 30-70 °C.

[0036] Particularly preferably, the dosage of the acid solution is 75-99% of the acid equivalent for reacting with all the fragmentation agent metals.

[0037] More preferably, the dosage of the acid solution is 80-85% of the acid equivalent for reacting with all the fragmentation agent metals.

[0038] Preferably, the step 3 is: completely oxidizing the fragmentation agent metal in the acid leaching residue by means of aerobic calcination.

[0039] More preferably, step 3 is specifically as follows: subject the acid-soluble slag to aerobic calcination, introduce compressed air at 1-3 L / min, heat it to 500-550 °C at a heating rate not higher than 5 °C / min, hold for 2-4 h, and then heat it to 700-750 °C and hold for 2-4 h.

[0040] Further preferably, after the calcined material after the aerobic calcination in step 3 is mixed evenly, perform phase analysis by XRD, then take a sample, dissolve it with acid, and detect whether there is hydrogen with a hydrogen gas alarm to determine whether the active fragmentation agent metal has been completely oxidized.

[0041] Preferably, the mixed material needs to be ball-milled between step 3 and step 4. After ball-milling, the mixed material needs to be sampled to detect the particle size distribution, and the particle size D100 of the mixed material is controlled to be less than 74 microns.

[0042] Further, if the noble metal is ruthenium, in step 4, the refining treatment method is as follows:

[0043] Subject the mixed material obtained in step 3 to alkali fusion and water leaching treatment in sequence. The water leaching solution obtained after water leaching is introduced with chlorine gas, and ruthenium tetroxide obtained by the reaction is collected by distillation. Convert ruthenium tetroxide into ammonium chlororuthenate, and obtain ruthenium powder after calcination reduction of ammonium chlororuthenate;

[0044] Further, if the noble metal is osmium, in step 4, the refining treatment method is as follows:

[0045] When the fragmentation agent metal in the acid leaching slag is subjected to aerobic calcination in step 3, part of the osmium will be oxidized into osmium tetroxide and volatilize. Absorb it with sodium hydroxide to obtain the first osmium-containing absorption solution. Subject the mixed material obtained in step 3 to alkali fusion and water leaching. The water leaching solution is introduced with chlorine gas and distilled, and then absorbed with sodium hydroxide to obtain the second absorption solution. Combine the two parts of the absorption solution, adjust the pH value by adding hydrochloric acid, use hydrazine hydrate to reduce to obtain osmium black, and obtain high-purity osmium powder after pickling, washing, drying, and hydrogen reduction;

[0046] Further, if the noble metal is iridium, in step 4, the refining treatment method is as follows:

[0047] Dissolve the mixed material obtained in step 3 to obtain an iridium-containing solution, then add ammonium chloride to obtain ammonium chloroiridate, further purify it by sulfidation, and finally obtain high-purity iridium powder after calcination reduction.

[0048] Specifically, the present invention also provides a method for recovering ruthenium metal from a ruthenium alloy, including the following steps:

[0049] Step 1: Mix and melt the ruthenium alloy and the fragmentation agent metal in sequence, and perform fragmentation treatment to obtain brittle alloy particles;

[0050] Step 2: Mix and react the brittle alloy particles with the acid solution to obtain a solution and acid leaching residue; the acid solution reacts with some of the fragmentation agent metals in the brittle alloy particles, and then solid-liquid separation is carried out to obtain a solution and acid leaching residue. The acid leaching residue contains at least some of the fragmentation agent metals in the form of elemental substances;

[0051] Step 3: Completely oxidize the fragmentation agent metals in the acid leaching residue to obtain a mixed material;

[0052] Step 4: Sequentially perform alkali fusion, water leaching and refining treatments on the mixed material obtained in Step 3 to obtain high-purity ruthenium powder.

[0053] Preferably, the fragmentation agent metal is aluminum or / and zinc.

[0054] Preferably, in Step 1, the weight ratio of the ruthenium alloy to the fragmentation agent metal is 1:1.1 - 1.4, the temperature of the mixed melting is 900 - 1100 °C, the heat preservation time is 30 - 60 min, and the size of the brittle alloy particles after fragmentation treatment in Step 1 is less than 10 mm.

[0055] Further preferably, in Step 1, the dosage of the fragmentation agent metal is at least 1.1 times, 1.2 times, 1.3 times, 1.4 times the weight of the alloy containing precious metals.

[0056] Preferably, the fragmentation agent metal is aluminum or / and zinc.

[0057] Furthermore, the fragmentation agent metal is preferably aluminum, zinc, but aluminum-zinc mixtures are not excluded, and the aluminum-zinc mixtures can be mixed in any proportion.

[0058] Furthermore, the maximum diameter of the added ruthenium alloy and fragmentation agent metal blocks does not exceed 180 mm, and the height does not exceed 330 mm.

[0059] Preferably, in Step 2, the fragmentation agent metal in the acid leaching residue in the form of elemental substances is equivalent to 1 wt% - 25 wt% of the total amount of the fragmentation agent metal.

[0060] Further preferably, in Step 2, the fragmentation agent metal in the acid leaching residue in the form of elemental substances is equivalent to 5 wt% - 20 wt% of the total amount of the fragmentation agent metal.

[0061] Further preferably, in Step 2, the fragmentation agent metal in the acid leaching residue in the form of elemental substances is equivalent to 7 wt% - 18 wt% of the total amount of the fragmentation agent metal.

[0062] Further preferably, in Step 2, the fragmentation agent metal in the acid leaching residue in the form of elemental substances is equivalent to 9 wt% - 16 wt% of the total amount of the fragmentation agent metal.

[0063] Further preferably, in the step 2, the fragmentation agent metal in the acid leaching residue in the form of a simple substance is equivalent to 11 wt% - 15 wt% of the total amount of the fragmentation agent metal.

[0064] Preferably, in the step 2, the fragmentation agent metal in the acid leaching residue in the form of a simple substance is equivalent to 5 wt%, 7 wt%, 9 wt%, 11 wt%, 13 wt%, 15 wt%, 17 wt%, 19 wt% of the total amount of the fragmentation agent metal.

[0065] Preferably, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:5 - 8; the acid in the acid solution is sulfuric acid or hydrochloric acid; the reaction time is 4 - 6 h, and the temperature is 20 - 90 °C.

[0066] Further, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:5.5 - 7.5.

[0067] Further, in the step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:6 - 7.

[0068] Further, in the step 2, the reaction time is 3 - 5 h, and the temperature is 30 - 70 °C.

[0069] Particularly preferably, the dosage of the acid solution is 75 - 99% of the acid equivalent for reacting with all the fragmentation agent metal.

[0070] Further preferably, the dosage of the acid solution is 80 - 85% of the acid equivalent for reacting with all the fragmentation agent metal.

[0071] Preferably, the step 3 is specifically as follows: subject the acid leaching residue to aerobic calcination, introduce compressed air at 1 - 3 L / min, heat it to 500 - 550 °C at a heating rate not higher than 5 °C / min, keep it warm for 2 - 4 h, and then heat it to 700 - 750 °C and keep it warm for 2 - 4 h.

[0072] More preferably, after the aerobic calcination in the step 3 is completed, the calcined material is mixed evenly and subjected to phase analysis by XRD, and then a sample is taken and dissolved in acid and detected by a hydrogen gas detector to determine whether the active fragmentation agent metal has been completely removed.

[0073] Further, the step 4 includes the following specific steps:

[0074] Step 41: Grind the calcined material obtained in the step 3 to obtain a ground material;

[0075] Step 42: Sequentially subject the ground material obtained in the step 41 to alkali fusion, water leaching, and centrifugation to obtain a water leaching solution and a water leaching residue;

[0076] Step 43: Refine the water leaching solution obtained in Step 42 to obtain high-purity ruthenium powder.

[0077] Furthermore, in the ball milling process of Step 41, zirconia beads need to be added, and the mass ratio of the calcined material to the zirconia beads is 1:1 - 3, the ball milling time is 8 - 12 h; the particle size D100 of the ball milled material after ball milling is less than 74 μm.

[0078] Preferably, Step 4 further includes the following steps:

[0079] Step 44: Dry the water leaching residue obtained in Step 42 and return it to Step 41 to sequentially carry out ball milling, alkali fusion, water leaching, centrifugation, and refining to obtain high-purity ruthenium powder.

[0080] Further, the specific operation of Step 42 is to add the ball milled material obtained in Step 41 to a mixture of potassium nitrate and potassium hydroxide, load it into the crucible of the heating furnace, heat it to 550 - 650 °C, then start stirring at a stirring speed of 5 - 10 r / min and keep it warm for 3 - 5 h, pour the melt into the stainless steel reaction kettle, add tap water to dissolve the cooled alkali fusion block according to a solid-liquid ratio of 1:6 - 15 for 2 - 4 h, and finally centrifuge to obtain the water leaching solution and the water leaching residue; the mass ratio of the ball milled material to potassium nitrate and potassium hydroxide is 1:1 - 2:3 - 4.

[0081] Furthermore, the water leaching solution and the water leaching residue obtained by centrifugation in Step 42 need to be sampled to detect the ruthenium content in the solution and calculate the leaching rate of ruthenium.

[0082] Further, the specific operation of Step 43 is as follows: Add the water leaching solution into the distillation kettle, start stirring, with a rotation speed of 100 - 150 r / min, then heat the solution temperature to 50 - 60 °C, and introduce chlorine gas for distillation. The flow rate of chlorine gas is 1 - 1.5 m 3 / h, and the chlorination time is 3 - 4 h. The distilled ruthenium tetroxide is absorbed by 6M - 12M analytical pure hydrochloric acid to obtain a ruthenium-containing absorption solution, and then ammonium ruthenium chloride is precipitated by adding analytical pure ammonium chloride according to the conventional method, and finally high-purity ruthenium powder is obtained through calcination and reduction.

[0083] Beneficial Effects

[0084] Compared with the prior art, the present invention has the following advantages:

[0085] (1) Retain part of the fragmentation agent metal during the stage of the reaction between the brittle alloy particles and the acid solution, ensuring that all the acid solution reacts with the fragmentation agent metal, avoiding excessive acid consumption, so that the highly active precious metal after fragmentation reacts with the acid, thus causing precious metal loss, improving the direct recovery rate of precious metals, and at the same time avoiding excessive acid consumption, so that a large amount of water is required to wash and neutralize the acid leaching residue after acid dissolution, reducing the amount of wastewater generated and the wastewater treatment cost;

[0086] (2) Use a two-step method to remove the fragmentation agent metal: First, only react with part of the fragmentation agent metal during the mixing and reaction stage of the brittle alloy particles and the acid solution, and then further oxidize the remaining fragmentation agent metal by means of aerobic calcination. At the same time, the acid gas is removed, avoiding the use of a large amount of acid and water for further acid leaching or washing the acid leaching residue, reducing the treatment cost, and greatly improving the direct recovery rate of ruthenium;

[0087] (3) Adopt a segmented heating method during the aerobic calcination stage. The heating in the first stage oxidizes all the fragmentation agent metal on the outside, and the second stage is to completely melt the fragmentation agent metal so that oxygen can fully contact and oxidize all the fragmentation agent metal; The segmented heating and heat preservation oxidize the fragmentation agent metal on the outer layer and the inner layer comprehensively, saving energy and being efficient.

[0088] (4) Control the weight ratio of the fragmentation agent metal and the precious metal alloy in step 1, reduce the chemical reagents required for later acid dissolution of the fragmentation agent metal and the amount of wastewater generated, reduce the increase in treatment costs, and avoid the phenomenon that the molten alloy is directly water quenched, resulting in excessive high-temperature active fragmentation agent metal liquid droplets dropping into the water and easily causing violent explosions. Specific embodiments

[0089] The following further describes the present invention in conjunction with embodiments, but does not constitute any limitation to the present invention. Any limited modifications made within the scope of the claims of the present invention are still within the scope of the claims of the present invention.

[0090] In order to elaborate on the technical content of the present invention, the following further explains in conjunction with the embodiments.

[0091] Example 1

[0092] A method for recovering ruthenium metal from a ruthenium alloy, comprising the following steps:

[0093] Step 1: Add the ruthenium alloy and aluminum blocks into a magnesia crucible according to a weight ratio of 1:1.2, then load it into an induction melting furnace, turn on the cooling water at 12 L / min, control the constant power at 14 kW so that the temperature of the mixed melting is 1000 °C, keep warm for 50 min, aluminum and the ruthenium alloy form a brittle material, and finally start the crucible rotation motor of the melting furnace to pour the melt into a dry copper disk for cooling;

[0094] Step 2: Put the cooled material obtained in step 1 into a hammer crusher and crush it into particles less than 10 mm;

[0095] Step 3: Add the particles obtained by crushing in Step 2 to an acid solution at a solid-liquid ratio of 1:6 for acid leaching. The acid added is hydrochloric acid, the acid consumption is 82% of the acid equivalent, the reaction time is 4 h, the temperature is 45 °C, and then solid-liquid separation is carried out to obtain acid leaching residues;

[0096] Step 4: Load the acid leaching residues obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 550 °C at a heating rate of 5 °C / min, and keep it warm for 2 h; then heat it to 720 °C and keep it warm for 3 h to obtain calcined materials;

[0097] Step 5: Add the calcined materials obtained in Step 4 to a polyurethane ball milling tank, add zirconia beads according to a mass ratio of calcined materials:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns;

[0098] Step 6: Mix the ball milled materials obtained in Step 5 according to a mass ratio of ball milled materials:potassium nitrate:potassium hydroxide = 1:2:3, load them into a crucible of a heating furnace, raise the temperature to 600 °C, start stirring at 10 r / min, keep it warm for 4 h, pour the melt into a stainless steel reaction kettle, add tap water to dissolve the cooled alkali fusion block at a solid-liquid ratio of 1:8 for 3 h, and finally centrifuge to obtain a water leaching solution and water leaching residues. The water leaching residues are dried and returned to ball milling, and the water leaching solution enters the next step of refining;

[0099] Step 7: Add 70 L of the water leaching solution to a 100 L distillation kettle, start stirring at a rotation speed of 120 r / min, then heat the solution temperature to 55 °C, introduce chlorine gas for distillation, the flow rate of chlorine gas is 1.5 m 3 / h, and the chlorination time is 3 h; the ruthenium tetroxide distilled out is absorbed with 10 M analytical pure hydrochloric acid to obtain a ruthenium-containing absorption solution, and then ammonium chloroplatinate is precipitated by adding analytical pure ammonium chloride according to the conventional method. Finally, high-purity ruthenium powder is obtained through calcination and reduction.

[0100] Example 2

[0101] A method for recovering ruthenium metal from a ruthenium alloy, comprising the following steps:

[0102] Step 1: Add the ruthenium alloy and aluminum blocks to a magnesia crucible at a weight ratio of 1:1.4, then load them into an induction melting furnace, turn on the cooling water at 15 L / min, control the constant power at 14 kW to make the temperature of the mixed melting 1100 °C, keep it warm for 60 min, aluminum and the ruthenium alloy form a brittle material, and finally start the crucible rotation motor of the melting furnace to pour the melt into a dry copper plate for cooling;

[0103] Step 2: Put the cooled materials obtained in Step 1 into a hammer crusher and crush them into particles smaller than 10 mm;

[0104] Step 3: Add the particles obtained by crushing in Step 2 to an acid solution at a solid-liquid ratio of 1:8 for acid dissolution. The acid added is hydrochloric acid, the acid consumption is 85% of the acid equivalent, the reaction time is 6 h, the temperature is 65 °C, and then solid-liquid separation is carried out to obtain acid leaching residue;

[0105] Step 4: Load the acid leaching residue obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 550 °C at a heating rate of 5 °C / min, and hold for 3 h; then heat it to 750 °C and hold for 4 h to obtain a calcined material;

[0106] Step 5: Add the calcined material obtained in Step 4 to a polyurethane ball mill tank, add zirconia beads at a mass ratio of calcined material:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns;

[0107] Step 6: Mix the ball milled material obtained in Step 5 with potassium nitrate and potassium hydroxide in a mass ratio of ball milled material:potassium nitrate:potassium hydroxide = 1:2:3, load it into a crucible of a heating furnace, raise the temperature to 650 °C, start stirring at 10 r / min, hold for 5 h, pour the melt into a stainless steel reaction kettle, dissolve the cooled alkali fusion block in tap water at a solid-liquid ratio of 1:15 for 4 h, and finally centrifuge to obtain a water leaching solution and a water leaching residue. The water leaching residue is dried and returned to the ball mill, and the water leaching solution enters the next step of refining;

[0108] Step 7: Add 70 L of the water leaching solution to a 100 L distillation kettle, start stirring at a rotation speed of 150 r / min, then heat the solution temperature to 60 °C, introduce chlorine gas for distillation, and the flow rate of chlorine gas is 1.5 m 3 / h, and the chlorination time is 4 h; the ruthenium tetroxide distilled out is absorbed by 12 M analytical pure hydrochloric acid to obtain a ruthenium-containing absorption solution, and then ammonium chloroplatinate is precipitated by adding analytical pure ammonium chloride according to the conventional method. Finally, high-purity ruthenium powder is obtained through calcination and reduction.

[0109] Example 3

[0110] A method for recovering ruthenium metal from a ruthenium alloy, comprising the following steps:

[0111] Step 1: Add the ruthenium alloy and aluminum block to a magnesia crucible at a weight ratio of 1:1.1, then load it into an induction melting furnace, turn on the cooling water at 10 L / min, and control the constant power at 14 kW to make the temperature of the mixed melting 900 °C, hold for 30 min, aluminum and the ruthenium alloy form a brittle material, and finally start the crucible rotation motor of the melting furnace to pour the melt into a dry copper plate for cooling;

[0112] Step 2: Put the cooled material obtained in Step 1 into a hammer crusher and crush it into particles less than 10 mm;

[0113] Step 3: Add the particles obtained by crushing in Step 2 to an acid solution at a solid-liquid ratio of 1:8 for acid leaching. The acid added is sulfuric acid, the acid consumption is 80% of the acid equivalent, the reaction time is 4 h, the temperature is 35 °C, and then solid-liquid separation is carried out to obtain acid leaching residue;

[0114] Step 4: Load the acid leaching residue obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 500 °C at a heating rate of 5 °C / min, and keep it warm for 2 h; then heat it to 700 °C and keep it warm for 2 h to obtain a calcined material;

[0115] Step 5: Add the calcined material obtained in Step 4 to a polyurethane ball mill tank, add zirconia beads at a mass ratio of calcined material:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns;

[0116] Step 6: Mix the ball milled material obtained in Step 5 with potassium nitrate and potassium hydroxide in a mass ratio of ball milled material:potassium nitrate:potassium hydroxide = 1:1:3, load it into a crucible of a heating furnace, heat it to 550 °C, start stirring at 5 r / min, keep it warm for 3 h, pour the melt into a stainless steel reaction kettle, add tap water to dissolve the cooled alkali fusion block at a solid-liquid ratio of 1:6 for 2 h, and finally centrifuge to obtain a water leaching solution and a water leaching residue. The water leaching residue is dried and returned to the ball mill, and the water leaching solution enters the next step of refining;

[0117] Step 7: Add 70 L of the water leaching solution to a 100 L distillation kettle, start stirring at a speed of 150 r / min, then heat the solution temperature to 50 °C, introduce chlorine gas for distillation, the flow rate of chlorine gas is 1 m 3 / h, and the chlorination time is 3 h; the ruthenium tetroxide distilled out is absorbed by 6 M analytical pure hydrochloric acid to obtain a ruthenium-containing absorption solution, and then ammonium chloroplatinate is precipitated by adding analytical pure ammonium chloride according to the conventional method. Finally, high-purity ruthenium powder is obtained through calcination and reduction.

[0118] Example 4

[0119] A method for recovering osmium metal from an osmium alloy, comprising the following steps:

[0120] Step 1: Add the osmium alloy and aluminum block to a magnesia crucible at a weight ratio of 1:1.2, then load it into an induction melting furnace, turn on the cooling water at 12 L / min, and control the constant power at 14 kW so that the temperature of the mixed melting is 1000 °C, keep it warm for 50 min, aluminum and the ruthenium alloy form a brittle material, and finally start the crucible rotation motor of the melting furnace, and pour the melt into a dry copper plate for cooling;

[0121] Step 2: Put the cooled material obtained in Step 1 into a hammer crusher and crush it into particles less than 10 mm;

[0122] Step 3: Add the particles obtained by crushing in Step 2 to an acid solution at a solid-liquid ratio of 1:6 for acid leaching. The acid added is hydrochloric acid, the acid consumption is 82% of the acid equivalent, the reaction time is 4 h, the temperature is 45 °C, and then solid-liquid separation is carried out to obtain acid leaching residue;

[0123] Step 4: Load the acid leaching residue obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 550 °C at a heating rate of 5 °C / min, and keep it warm for 2 h; then heat it to 720 °C and keep it warm for 3 h to obtain a calcined material. Pass the calcined tail gas into a 25% analytical pure sodium hydroxide solution to obtain the first osmium-containing absorption solution;

[0124] Step 6: Put the calcined material obtained in Step 4 into a polyurethane ball mill pot, add zirconia beads according to the mass ratio of calcined material:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns;

[0125] Step 9: Mix the ball milled material obtained in Step 5 according to the mass ratio of ball milled material:sodium peroxide:sodium hydroxide = 1:2:0.5, load it into the crucible of the heating furnace, raise the temperature to 600 °C, start stirring at 10 r / min, keep it warm for 4 h, pour the melt into a stainless steel reaction kettle, add tap water to dissolve the cooled alkali fusion block at a solid-liquid ratio of 1:8 for 3 h, and finally centrifuge to obtain a water leaching solution and a water leaching residue. The water leaching residue is dried and returned to the ball mill, and the water leaching solution enters the next refining step;

[0126] Step 7: Add 70 L of the water leaching solution to a 100 L distillation kettle, start stirring at a rotation speed of 120 r / min, then heat the solution temperature to 55 °C, introduce chlorine gas for distillation, and the flow rate of chlorine gas is 1.5 m 3 / h, and the chlorination time is 3 h; the distilled osmium tetroxide is absorbed by a 25% analytical pure sodium hydroxide solution to obtain the second osmium-containing absorption solution;

[0127] Step 8: Combine the absorption solutions in Step 4 and Step 7, add hydrochloric acid to adjust the pH to 8, add hydrazine hydrate for reduction to obtain osmium black, and obtain high-purity osmium powder through pickling, washing, drying, and hydrogen reduction.

[0128] Example 5

[0129] A method for recovering iridium metal from an iridium alloy, comprising the following steps:

[0130] Step 1: Add the iridium alloy and aluminum block to a magnesia crucible at a weight ratio of 1:1.2, then load it into an induction melting furnace, turn on the cooling water at 12 L / min, control the constant power at 14 kW to make the temperature of the mixed melting 1000 °C, keep it warm for 50 min, and aluminum and the ruthenium alloy form a brittle material. Finally, start the crucible rotation motor of the melting furnace and pour the melt into a dry copper plate for cooling;

[0131] Step 2: Put the cooled material obtained in Step 1 into a hammer crusher and crush it into particles smaller than 10 mm.

[0132] Step 3: Add the particles obtained by crushing in Step 2 to an acid solution according to a solid-liquid ratio of 1:6. The acid added is hydrochloric acid, the amount of acid used is 82% of the acid equivalent, the reaction time is 4 h, the temperature is 45 °C, and then solid-liquid separation is carried out to obtain acid leaching residues.

[0133] Step 4: Load the acid leaching residues obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 550 °C at a heating rate of 5 °C / min, and keep it warm for 2 h; then heat it to 720 °C and keep it warm for 3 h to obtain calcined materials.

[0134] Step 5: Put the calcined materials obtained in Step 4 into a polyurethane ball milling tank, add zirconia beads according to a mass ratio of calcined materials:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns.

[0135] Step 6: Mix the ball milled materials obtained in Step 5 according to a mass ratio of ball milled materials:sodium peroxide:sodium hydroxide of 1:3:1, load them into a crucible of a heating furnace, raise the temperature to 600 °C, start stirring at 10 r / min, keep it warm for 4 h, pour the melt into a stainless steel reaction kettle, add tap water to dissolve the cooled alkali fusion block according to a solid-liquid ratio of 1:8 for 3 h, centrifuge, collect the water leaching solution, fully dissolve the water leaching residues with aqua regia, filter and wash to obtain a solution of iridium chloroacid.

[0136] Step 7: Add excessive ammonium chloride to the iridium chloroacid solution and centrifuge to obtain a black precipitate of (NH4)2IrCl6.

[0137] Step 8: Pulverize the (NH4)2IrCl6 precipitate with deionized water, then add an appropriate amount of hydrazine hydrate, heat and stir the reaction until it is completely converted into a solution of (NH4)3IrCl6, and filter; add a dilute solution of ammonium sulfide to the solution of ammonium iridium chloroacid at room temperature, let it stand, and filter; then heat the filtrate, slowly add ammonium sulfide solution for secondary sulfidation, let it stand, and filter to obtain a pure solution of (NH4)3IrCl6.

[0138] Step 9: Add hydrochloric acid and hydrogen peroxide to the pure solution of (NH4)3IrCl6 for oxidation precipitation, centrifuge to obtain a pure precipitate of (NH4)2IrCl6, and finally obtain high-purity iridium powder through calcination and reduction.

[0139] Example 6

[0140] A method for recovering ruthenium, osmium, and iridium metals from an alloy containing ruthenium, osmium, and iridium, comprising the following steps:

[0141] Step 1: Add ruthenium, osmium, iridium alloy and aluminum block into a magnesia crucible at a weight ratio of 1:1.2, then load it into an induction melting furnace. Turn on the cooling water at 12 L / min and control the constant power at 14 kW to make the temperature of the mixed melting reach 1000 °C. Keep the temperature for 50 min. Aluminum and ruthenium alloy form a brittle material. Finally, start the crucible rotation motor of the melting furnace and pour the melt into a dry copper disk for cooling;

[0142] Step 2: Put the cooled material obtained in Step 1 into a hammer crusher and crush it into particles smaller than 10 mm;

[0143] Step 3: Add the particles obtained by crushing in Step 2 into an acid solution at a solid-liquid ratio of 1:6. The acid added is hydrochloric acid, and the acid consumption is 82% of the acid equivalent. The reaction time is 4 h and the temperature is 45 °C. Then carry out solid-liquid separation to obtain acid leaching residue;

[0144] Step 4: Load the acid leaching residue obtained in Step 3 into a tubular atmosphere furnace, introduce compressed air at 2 L / min, heat it to 550 °C at a heating rate of 5 °C / min, and keep the temperature for 2 h; then heat it to 720 °C and keep the temperature for 3 h to obtain a calcined material. Pass the calcined tail gas into a 25% analytical pure sodium hydroxide solution to obtain the first osmium-containing absorption solution;

[0145] Step 5: Put the calcined material obtained in Step 4 into a polyurethane ball mill tank, add zirconia beads at a mass ratio of calcined material:zirconia beads of 1:2, and ball mill until the particle size D100 of all calcined materials is less than 74 microns;

[0146] Step 6: Mix the ball milled material obtained in Step 5 according to the mass ratio of ball milled material:sodium peroxide:sodium hydroxide = 1:3:2, load it into the crucible of a heating furnace, raise the temperature to 600 °C, turn on the stirring at 10 r / min, keep the temperature for 4 h, pour the melt into a stainless steel reaction kettle, and dissolve the cooled alkali fusion block in tap water at a solid-liquid ratio of 1:8 for 3 h. Finally, centrifuge to obtain iridium-containing water leaching residue and ruthenium- and osmium-containing water leaching solution;

[0147] Step 7: Dissolve the iridium-containing water leaching residue thoroughly with aqua regia, filter and wash to obtain a chloroiridic acid solution; add excessive ammonium chloride to the chloroiridic acid solution and centrifuge to obtain a black (NH4)2IrCl6 precipitate;

[0148] Step 8: Pulverize the (NH4)2IrCl6 precipitate with deionized water, then add an appropriate amount of hydrazine hydrate, heat and stir the reaction until it is completely converted into a (NH4)3IrCl6 solution, and filter; at room temperature, add a dilute ammonium sulfide solution to the ammonium chloroiridate solution, let it stand, and filter; then heat the filtrate, slowly add an ammonium sulfide solution for secondary sulfidation, let it stand, and filter to obtain a pure (NH4)3IrCl6 solution;

[0149] Step 9: Add hydrochloric acid and hydrogen peroxide to the pure (NH4)3IrCl6 solution for oxidation precipitation, and centrifuge to obtain pure (NH4)2IrCl6 precipitate. Finally, high-purity iridium powder is obtained through calcination and reduction.

[0150] Step 10: Add 70 L of the water leaching solution containing ruthenium and osmium to a 100 L distillation kettle, start stirring at a speed of 120 r / min, then heat the solution temperature to 55 °C, and introduce chlorine for distillation. The flow rate of chlorine is 1.5 m 3 / h, and the chlorination time is 3 h; distill ruthenium and osmium in the solution into the gas phase, and introduce the gas into an analytical pure hydrochloric acid solution to obtain a hydrochloric acid solution containing ruthenium; then introduce the gas after separating ruthenium into a 25% analytical pure sodium hydroxide solution to obtain a second osmium-containing absorption solution.

[0151] Step 11: Add analytical pure ammonium chloride to the ruthenium-containing hydrochloric acid absorption solution obtained in Step 10 to precipitate ammonium chlororuthenate, and finally obtain high-purity ruthenium powder through calcination and reduction.

[0152] Step 12: Combine the two parts of the osmium-containing absorption solutions from Step 4 and Step 10, adjust the pH to 8 with hydrochloric acid, add hydrazine hydrate for reduction to obtain osmium black, and obtain high-purity osmium powder through pickling, washing, drying, and hydrogen reduction.

[0153] Comparative Example 1

[0154] Basically the same as Example 1, except that the acid consumption in Step 3 is 200% of the acid equivalent for the reaction of the acid solution with all the fragmentation agent metals; all the fragmentation agent metals in the acid leaching residue are dissolved, and the solid after acid leaching is washed with water multiple times until neutral.

[0155] Effect Test

[0156] Calculation method: 1. Direct recovery rate = (mass of precious metals in the recovered product / mass of precious metals in the raw material alloy) * 100%; 2. Waste water volume generated per kilogram of precious metal alloy treated = volume of waste water generated in batches / mass of precious metal alloy treated in batches.

[0157] Perform precious metal direct recovery rate tests on Examples 1-6 and Comparative Example 1, and the test results are shown in Table 1.

[0158] Table 1

[0159]

[0160]

[0161] It can be seen from the comparison between Example 1 and Comparative Example 1 that when too much acid solution is used to treat aluminum with strong reaction activity, part of the ruthenium metal will also react with the acid solution, thus affecting the direct recovery rate of ruthenium, and the amount of wastewater generated per unit mass of ruthenium metal alloy treated increases significantly;

[0162] Therefore, the technical solution of the present invention effectively removes the fragmented metal with strong reaction activity, further improves the direct recovery rate of precious metals, avoids the loss of precious metals, reduces the amount of wastewater, and effectively controls the wastewater treatment cost.

[0163] The embodiments presented herein are only the implementation manners selected according to the combinations of all possible embodiments. The appended claims should not be limited by the embodiments illustrating the present invention. Some numerical ranges used in the claims include sub-ranges within them, and the variations within these ranges should also be covered by the appended claims.

Claims

1. A method for recovering precious metals from a precious metal alloy, characterized in that, It includes the following steps: Step 1: Mix and melt an alloy containing precious metals and a fragmentation agent metal, and perform fragmentation treatment to obtain brittle alloy particles; Step 2: Mix and react the brittle alloy particles with an acid solution to obtain a solution and an acid leaching residue; the reaction activity of the fragmentation agent metal in the brittle alloy particles with the acid is higher than that of the precious metal in the brittle alloy particles with the acid; at least part of the fragmentation agent metal in the acid leaching residue exists in the form of a single substance; Step 3: Completely oxidize the fragmentation agent metal in the acid leaching residue to obtain a mixed material; Step 4: Refine the mixed material to extract a precious metal compound or a precious metal single substance; The precious metal is at least one of ruthenium, osmium, and iridium; In Step 1, the dosage of the fragmentation agent metal is at least 1.1 - 1.4 times the weight of the alloy containing precious metals; the fragmentation agent metal is aluminum or / and zinc; In Step 2, the dosage of the acid solution is 75 - 99% of the acid equivalent for reacting with all the fragmentation agent metals.

2. The method for recovering precious metals from a precious metal alloy according to claim 1, characterized in that, In Step 2, the fragmentation agent metal in the acid leaching residue existing in the form of a single substance is equivalent to 1wt% - 25wt% of the total amount of the fragmentation agent metal.

3. The method for recovering precious metals from a precious metal alloy according to claim 2, characterized in that, In Step 2, the fragmentation agent metal in the acid leaching residue existing in the form of a single substance is equivalent to 5wt% - 20wt% of the total amount of the fragmentation agent metal.

4. The method for recovering precious metals from a precious metal alloy according to claim 1, characterized in that, In Step 2, the mass ratio of the brittle alloy particles to the acid solution is 1:5 - 8; the acid in the acid solution is sulfuric acid or hydrochloric acid; the reaction time is 4 - 6h, and the temperature is 20 - 90°C.

5. The method for recovering precious metals from a precious metal alloy according to claim 1, characterized in that, Step 3 is: Completely oxidize the fragmentation agent metal in the acid leaching residue by means of aerobic calcination.

6. The method for recovering precious metals from a precious metal alloy according to claim 5, characterized in that, Step 3 specifically is: Aerobically calcine the acid leaching residue, introduce compressed air at 1 - 3L / min, heat it to 500 - 550°C at a heating rate not higher than 5°C / min, keep it warm for 2 - 4h, and then heat it to 700 - 750°C and keep it warm for 2 - 4h.

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