Trapping and recycling method for ruthenium in ruthenium-containing waste

Abstract

The invention belongs to the technical field of precious metal enrichment and recovery, and discloses a method for capturing and recovering ruthenium in ruthenium-containing waste, which comprises the following steps: step 1, sequentially carrying out acid dissolution and impurity removal on the ruthenium-containing waste, and separating to obtain ruthenium-containing filter residues; 2, a trapping agent, a slag former and a reducing agent are added into the ruthenium-containing filter residues, and then matte smelting is conducted to obtain smelting residues and alloy; 3, the alloy is sequentially subjected to activating treatment, acid dissolution impurity removal and calcination, and ruthenium powder is obtained; wherein the slag former comprises calcium oxide and / or calcium carbonate, silicon dioxide, borax and sodium carbonate, and the dosage ratio of calcium element to boron element in the slag former is CaO: B2O3 = (1-3): 1 by mass. According to the method, the calcium-boron ratio in the slag former is controlled in the ruthenium trapping and recycling process through the matte smelting method, the separation effect of the slag phase and the alloy phase can be effectively improved, the ruthenium trapping effect can be improved, and the ruthenium recycling rate is increased.

Description

Technical Field

[0001] The invention relates to the technical field of precious metal enrichment and recovery, and in particular to a method for capturing and recovering ruthenium in ruthenium-containing waste materials. Background Art

[0002] Ruthenium, a platinum group metal, has strong catalytic properties and high stability. In the field of semiconductor integrated circuit industry, magnetron sputtering is used to sputter ruthenium target medium onto electronic devices to form a thin film medium to improve the performance of disks. During the sputtering spraying process, some precious metals are lost on the surface of parts (gaskets, baffles, etc.). The parts are mainly made of hard alloys mainly composed of ruthenium, chromium, and cobalt. After sputtering, the ruthenium content on the surface of the parts is about 1000-450000ppm (0.1-45wt%). In order to ensure the recycling of resources, this type of ruthenium-containing alloy waste is collected and recycled in the industry to prepare target-grade ruthenium powder again. Therefore, recycling ruthenium metal in sputtering spraying equipment is conducive to solving the problem of shortage of ruthenium target materials in industry, which has important economic significance.

[0003] Ruthenium is a precious metal that is difficult to dissolve. Its solubility is a problem in the resource recovery process. At the same time, the parts of sputtering target equipment are often hard alloys that are resistant to high temperatures and have high hardness. It is impossible to break the particles by conventional physical methods to recover and enrich ruthenium metal. Therefore, the industry often uses base metal activation to reduce the eutectic point and hardness of the alloy, so that the hard alloy is pulverized, and the base metal impurities are removed by acid leaching before the ruthenium metal recovery process is carried out.

[0004] Common methods for recovering ruthenium include oxidative distillation, chlorination, molten water leaching, alkaline hot pressure reduction, microwave leaching and extraction, etc.

[0005] For example, prior art 1 uses a combination of oxidative distillation and microwave leaching extraction to recover ruthenium. Prior art 1: Chinese patent application 201910743605.2 discloses a method for recovering ruthenium trichloride in ruthenium plating waste liquid, comprising: S1: oxidative distillation of ruthenium plating waste liquid under acidic medium to obtain ruthenium trichloride and residual liquid; S2: extracting and back-extracting the residual liquid to obtain an aqueous phase; S3: oxidative distillation of the aqueous phase.

[0006] Just like the prior art 2, the chlorination method is used to recover ruthenium. Prior art 2: Chinese patent application 202310779867.0 A method for recovering ruthenium chloride from ruthenium-containing waste catalyst, the ruthenium-containing waste catalyst is roasted to remove organic matter mixed in the waste catalyst, and then ground after cooling to obtain ruthenium ash powder, the ruthenium ash powder is slowly added to the alkaline solution and stirred, and sodium hypochlorite solution is added to the mixed solution. After the sodium hypochlorite solution is added, the mixture is kept warm for reaction, and filtered to obtain a ruthenium-containing filtrate; formic acid is added to the ruthenium-containing filtrate, and a ruthenium hydroxide precipitate is obtained by reaction, and the precipitate is washed; the ruthenium hydroxide precipitate is added to a mixed solution composed of hydrochloric acid and hydroxylamine hydrochloride, the precipitate is dissolved to obtain a ruthenium trichloride solution, and the ruthenium trichloride is obtained by drying.

[0007] Both prior art 1 and prior art 2 use common methods to recover ruthenium, and both show good ruthenium recovery effects. However, the above methods are not very applicable to complex materials (i.e., low-grade materials) with a ruthenium content of 0.01-10wt%. The main common problems are lengthy reaction process, low direct ruthenium yield, high energy consumption and material consumption, and inability to process large amounts of materials.

[0008] Based on this, the technical problem to be solved by the present invention is: how to find a method for recovering ruthenium from complex materials with a ruthenium content of 0.01-10wt%, avoiding the problem of lengthy reaction process and at the same time improving the direct recovery rate of ruthenium. Summary of the invention

[0009] One of the purposes of the present invention is to provide a method for capturing and recovering ruthenium in ruthenium-containing waste. The ruthenium metal in the ruthenium-containing waste is captured and recovered by a matte making and smelting method, and low-grade ruthenium in the smelting system can be captured and recovered. It has the advantages of high direct recovery and short reaction process for complex materials with a ruthenium content of 0.01-10wt%.

[0010] To achieve the above object, the present invention provides a method for capturing and recovering ruthenium in ruthenium-containing waste, comprising the following steps:

[0011] Step 1: acid dissolving and removing impurities from the ruthenium-containing waste material in sequence, and separating to obtain a ruthenium-containing filter residue;

[0012] Step 2: adding a collector, a slag-making agent and a reducing agent to the ruthenium-containing filter residue, and then performing matte smelting to obtain smelting slag and a cobalt alloy;

[0013] Step 3: subjecting the cobalt alloy to activation treatment, acid dissolution and impurity removal, and calcination in sequence to obtain ruthenium powder;

[0014] Wherein, the slag-forming agent includes calcium oxide and / or calcium carbonate, silicon dioxide, borax, and soda ash, and the usage ratio of calcium element to boron element in the slag-forming agent is CaO:B2O3=1-3:1 by mass.

[0015] Based on the capture and recovery of ruthenium by matte smelting method, the present invention studies the ratio of boron and calcium in the slag-forming agent, and finds that the usage ratio of calcium and boron in the slag-forming agent (hereinafter referred to as calcium-boron ratio) not only affects the separation of slag phase and serpentine alloy phase in the smelting system, but also affects the capture efficiency of ruthenium. The reasons may be:

[0016] (1) The calcium-boron ratio affects the growth efficiency of calcium borate and the formation of the final slag phase, thereby affecting the separation effect of the slag phase and the cobalt alloy;

[0017] (2) The calcium-boron ratio affects the activation efficiency of calcium, thereby affecting the internal driving force of ruthenium ions during their release, and ultimately affecting the recovery rate of ruthenium.

[0018] Furthermore, the dosage of the slag-forming agent is 0.8-2 times the mass of the ruthenium-containing filter residue.

[0019] Preferably, the amount of calcium oxide and / or calcium carbonate, silicon dioxide, borax and soda ash in the slagging agent needs to meet the following conditions:

[0020] The amount ratio of boron to silicon (hereinafter referred to as boron-silicon ratio) is B2O3:SiO2=0.2-0.6:1 by mass, the amount ratio of calcium to silicon (hereinafter referred to as calcium-silicon ratio) is CaO:SiO2=0.5-1.2:1 by mass, the basicity is 0.9-1.5, the theoretical viscosity of the melt is 0.5-1.1, and the theoretical surface tension of the melt is 0.1-0.5.

[0021] On the basis of controlling the calcium-boron ratio, the present invention further controls the boron-silicon ratio, calcium-silicon ratio, alkalinity, viscosity and surface tension of each component of the slag-forming agent to itself or to the melt; wherein the combination of the boron-silicon ratio and the calcium-boron ratio can effectively control the slag phase to be generated in the form of glass slag that is incompatible with the matte alloy, that is, silicates and borates whose basic elements are [SiO3] and [BO3]; and the calcium-silicon ratio can inhibit the formation of SiO4 - The formation of Si x O y n- The long-chain glass network is polymerized and decomposed, and the melt viscosity is reduced; the control of alkalinity can provide the appropriate O for the entire melting system. 2- The amount of ruthenium metal in the slag is controlled to promote the reaction between the slag and the ruthenium metal; the control of the theoretical viscosity of the melt affects the fluidity of the ruthenium metal, which also affects the separation effect of the final slag phase and the silicate alloy; the control of the theoretical surface tension of the melt can avoid the generation of bubbles in the slag phase and reduce the inclusion of ruthenium metal in the slag phase. The present invention controls the boron-silicon ratio, calcium-silicon ratio, alkalinity, theoretical viscosity of the melt, and theoretical surface tension of the melt in the slag-making agent. The combination of the five produces a synergistic effect of improving the capture efficiency of ruthenium, and synergistically controls the formation of the slag phase and the silicate alloy phase, effectively improving the separation effect of the slag phase and the silicate alloy.

[0022] It should be noted that the basicity can be calculated based on the aluminum oxide in the ruthenium-containing material acting as an acidic oxide, combined with the slag-making agent, reducing agent, and trapping agent. The formula is:

[0023]

[0024] Where: m is mass; basicity refers to the ratio of the sum of the main basic oxide contents in the slag to the sum of the main acidic oxide contents, generally represented by R.

[0025] The theoretical melt viscosity is calculated based on the NPL model, which uses optical basicity to represent structural and cationic effects.

[0026] The optical basicity is Λ=∑ i (X i Λ i )

[0027] Λ i is the optical basicity of the ith oxide;

[0028] The theoretical values ​​of optical basicity of the oxides in the present invention are: Na2O=1.15, CaO=1.0, Al2O3=0.60, SiO2=0.48, B2O3=0.43;

[0029] X i is the mole fraction of cations in the oxide, that is, the mole fraction of oxygen atoms in the oxide in the slag, calculated as:

[0030] x i is the mole fraction of the ith oxide in the slag, n (o)i is the number of oxygen atoms in the oxide molecule.

[0031] The viscosity calculation formula of the NPL model is: Inη=InA+B / T

[0032] Where T is the temperature, parameter

[0033] Parameter A = exp[-232.69*Λ 2 +357.32*Λ-144.17]

[0034] The unit of parameter A is Pa.s, the unit of parameter B is K, and the unit of viscosity η is Pa.s.

[0035] Under the smelting condition of 1200-1400℃, the slag type is adjusted by adding slag-forming agent, reducing agent and collecting agent. When the theoretical viscosity is controlled at 0.5-1.1, the melt fluidity is the best and the interface between glass slag and matte alloy is clear.

[0036] The theoretical surface tension of the melt is calculated based on the surface tension factors of Na, Ca, Al, and Si when the aluminum oxide in the ruthenium-containing material acts as an acidic oxide and is combined with the slag-making agent. The calculation method is: σ=∑ i X i *σ i

[0037] Among them, X i is the mole fraction of the oxide in the ith i It is the surface tension factor, which is mainly involved in the calculation of the surface tension factor of glass slag of oxides at 1400℃ (unit N / m): Na2O=0.30, CaO=0.51, Al2O3=0.58, SiO2=0.29.

[0038] Preferably, the ruthenium-containing waste in step 1 is activated ruthenium-containing waste, and the activation step of the ruthenium-containing waste is: the ruthenium-containing waste and the aluminum block are configured in a mass ratio of 1:2-4, and activated at a temperature of 1000-1400°C for 60-90 minutes.

[0039] The activation step uses the aluminothermic activation method in the prior art to fully activate the ruthenium-containing waste, which can improve the reaction efficiency of acid dissolution and impurity removal and avoid the problem of lengthy reaction process.

[0040] Furthermore, the specific operation of step 1 is: adding 6-12 mol / L acid solution to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10-12ml to carry out acid dissolution and impurity removal at a temperature of 90-100°C, and then performing solid-liquid separation to obtain a ruthenium-containing filter residue.

[0041] Preferably, the acid solution is hydrochloric acid or sulfuric acid solution.

[0042] Furthermore, in order to ensure that the acid dissolution and impurity removal are complete, an existing hydrogen concentration detection instrument such as a hydrogen alarm or a hydrogen detector can be used to detect whether hydrogen is still generated in the acid melting system. If no hydrogen is generated, the acid dissolution and impurity removal step is completed and solid-liquid separation can be performed.

[0043] Furthermore, the specific operation of step 2 is: the ruthenium-containing filter residue is stirred and mixed evenly with a slag-forming agent and a reducing agent and put into a crucible, and then a collector is added and spread on top, and the crucible is placed in a smelting furnace and smelted at a temperature of 1200-1400°C for 60-90 minutes, and then cast and cooled to obtain smelting slag and ruthenium alloy; wherein the amount of the reducing agent added is 20-30wt% of the ruthenium-containing filter residue, and the amount of the collector added is 30-50wt% of the ruthenium-containing filter residue.

[0044] Preferably, the collector is ferrous sulfide and the reducing agent is activated carbon.

[0045] Furthermore, the specific operation of the activation treatment in step 3 is: the cobalt alloy and the aluminum block are configured in a mass ratio of 1:2-4, and activated at a temperature of 1000-1400°C for 60-90 minutes to obtain the thermite alloy.

[0046] Furthermore, the specific operation of acid dissolution and impurity removal in step 3 is: adding 6-12 mol / L acid solution to the thermite alloy according to a solid-liquid ratio of 1g:10-12ml to carry out acid dissolution and impurity removal at a temperature of 90-100°C, and then separating the solid and liquid to obtain acid leaching solution and acid leaching residue.

[0047] After testing, the acid leaching solution still contains some ruthenium. Therefore, in order to avoid the loss of ruthenium and reduce the difficulty of sewage treatment, the step 3 further includes reduction precipitation. The specific operation of reduction precipitation is: adding hydrochloric acid or sodium hydroxide to the acid leaching solution to adjust the pH of the acid leaching solution to 4-5, adding a sulfide solution with a sulfide content of 20wt% to adjust the pH to 7-9, stirring at a constant temperature of 40-80°C for 50-60min, and separating the solid and liquid to obtain ruthenium-containing slag and waste liquid.

[0048] Preferably, the sulfide is at least one of sodium sulfide, zinc sulfide and magnesium sulfide.

[0049] Furthermore, the ruthenium-containing slag can be collected and then collected and recovered again in step 2 together with the ruthenium-containing filter residue obtained in step 1, and the waste liquid can be collected and handed over to a sewage treatment workshop for centralized treatment.

[0050] Furthermore, the specific operation of calcination in step 3 is: putting the acid leaching residue into a tubular furnace, introducing hydrogen at a gas flow rate of 3-5 L / min, and heating the temperature to 400-600° C. at a heating rate of 5-10° C. / min and keeping the temperature for 4-6 hours to obtain ruthenium powder.

[0051] Beneficial Effects

[0052] Compared with the prior art, the present invention has at least the following advantages:

[0053] (1) The present invention controls the calcium-boron ratio in the slag-forming agent during the process of capturing and recovering ruthenium by the matte smelting method, which can effectively improve the separation effect of the slag phase and the ruthenium alloy phase, and can also improve the capture effect of ruthenium and improve the recovery rate of ruthenium;

[0054] (2) The present invention controls the boron-silicon ratio, calcium-silicon ratio, alkalinity, melt theoretical viscosity, and melt theoretical surface tension in the slag-forming agent. The combination of the five factors produces a synergistic effect of improving the capture efficiency of ruthenium, and synergistically controls the formation of the slag phase and the cobalt alloy phase, thereby effectively improving the separation effect of the slag phase and the cobalt alloy. BRIEF DESCRIPTION OF THE DRAWINGS

[0055] The present invention is further described below in conjunction with the accompanying drawings and embodiments;

[0056] Figure 1 is a photo of the cobalt alloy of Example 1 of the present invention;

[0057] Figure 2 This is a photo of the smelting slag hanging on the wall in Comparative Example 3 of the present invention;

[0058] Figure 3 is a photo of the cobalt alloy of Comparative Example 3 of the present invention;

[0059] Figure 4 This is a photo of the copper alloy of Comparative Example 4 of the present invention. DETAILED DESCRIPTION

[0060] The present invention will be further described below in conjunction with the embodiments, but it does not constitute any limitation to the present invention. Any limited number of modifications made within the scope of the claims of the present invention are still within the scope of the claims of the present invention.

[0061] In order to explain the technical content of the present invention in detail, further description is given below in conjunction with the implementation modes.

[0062] In the following examples and comparative examples, the ruthenium content was measured by ICP.

[0063] Example 1

[0064] A method for capturing and recovering ruthenium from ruthenium-containing waste comprises the following steps:

[0065] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0066] Step 2: The ruthenium-containing filter residue, 31g borax, 48g sodium carbonate, 32g calcium oxide, 32g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 0.1g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0067] Calcium-boron ratio CaO: B2O3 = 2.86: 1, boron-silicon ratio B2O3: SiO2 = 0.35: 1, calcium-silicon ratio CaO: SiO2 = 1: 1, basicity = 1.50, melt theoretical viscosity = 0.77, melt theoretical surface tension = 0.34;

[0068] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0069] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 4.1 g.

[0070] Example 2

[0071] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0072] Step 2: The ruthenium-containing filter residue, 24g borax, 45g sodium carbonate, 8.75g calcium oxide, 15g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 0.13g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0073] Calcium-boron ratio CaO: B2O3 = 1:1, boron-silicon ratio B2O3: SiO2 = 0.58:1, calcium-silicon ratio CaO: SiO2 = 0.58:1, basicity = 1.36, melt theoretical viscosity = 0.84, melt theoretical surface tension = 0.31;

[0074] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0075] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 4.01g.

[0076] Example 3

[0077] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0078] Step 2: The ruthenium-containing filter residue, 28g borax, 45g sodium carbonate, 32.8g calcium oxide, 28g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 0.2g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0079] Calcium-boron ratio CaO: B2O3 = 3:1, boron-silicon ratio B2O3: SiO2 = 0.39:1, calcium-silicon ratio CaO: SiO2 = 1.17:1, basicity = 1.46, melt theoretical viscosity = 0.89, melt theoretical surface tension = 0.35;

[0080] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0081] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 4g.

[0082] Example 4

[0083] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0084] Step 2: The ruthenium-containing filter residue, 80.66g borax, 80g sodium carbonate, 5g calcium oxide, 10g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is placed in a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 0.22g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0085] Calcium-boron ratio CaO: B2O3 = 2.94: 1, boron-silicon ratio B2O3: SiO2 = 0.17: 1, calcium-silicon ratio CaO: SiO2 = 0.5: 1, basicity = 1.46, melt theoretical viscosity = 1.07, melt theoretical surface tension = 0.24;

[0086] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0087] Step 4: The acid leaching residue is placed in a tubular furnace, hydrogen is introduced at a gas flow rate of 4 L / min, and the temperature is increased to 600°C at a heating rate of 6°C / min and kept warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 3.9 g.

[0088] Comparative Example 1

[0089] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0090] Step 2: The ruthenium-containing filter residue, 24.7g borax, 31.5g sodium carbonate, 45g calcium oxide, 45g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 0.57g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0091] Calcium-boron ratio CaO: B2O3 = 5:1, boron-silicon ratio B2O3: SiO2 = 0.2:1, calcium-silicon ratio CaO: SiO2 = 1:1, basicity = 1.14, melt theoretical viscosity = 1.67, melt theoretical surface tension = 0.37;

[0092] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0093] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 3.2 g.

[0094] Comparative Example 2

[0095] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml at a temperature of 95°C±2°C for acid dissolution and impurity removal, and then solid-liquid separation is performed to obtain 100 g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 4.39 g after testing;

[0096] Step 2: The ruthenium-containing filter residue, 68.59g borax, 50g sodium carbonate, 25g calcium oxide, 30g silicon dioxide and 25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 34g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 2.9g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0097] Calcium-boron ratio CaO: B2O3 = 0.83: 1, boron-silicon ratio B2O3: SiO2 = 0.6: 1, calcium-silicon ratio CaO: SiO2 = 0.5: 1, basicity = 0.92, melt theoretical viscosity = 3.24, melt theoretical surface tension = 0.28;

[0098] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0099] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 1.2g.

[0100] Comparative Example 3

[0101] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml to carry out acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain 240g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 12.52g after testing;

[0102] Step 2: The ruthenium-containing filter residue, 39.64g borax, 5g sodium carbonate, 81.6g calcium oxide, 85g silicon dioxide and 60g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 115g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 5.7g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0103] Calcium-boron ratio CaO: B2O3 = 5.64: 1, boron-silicon ratio B2O3: SiO2 = 0.17: 1, calcium-silicon ratio CaO: SiO2 = 0.96: 1, basicity = 0.77, melt theoretical viscosity = 4.08, melt theoretical surface tension = 0.38;

[0104] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0105] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 2.6 g.

[0106] Comparative Example 4

[0107] Step 1: The ruthenium-containing waste material and the aluminum block are prepared in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes, and then a 10 mol / L acid solution is added to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10ml to carry out acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain 121g of ruthenium-containing filter residue. The ruthenium content in the ruthenium-containing filter residue is 5.32g after testing;

[0108] Step 2: The ruthenium-containing filter residue, 32g borax, 20g sodium carbonate, 31g calcium oxide, 50g silicon dioxide and 30.25g activated carbon as a reducing agent are stirred and evenly mixed and put into a crucible, and then 58g ferrous sulfide is added and spread on top, and the crucible is put into a smelting furnace and smelted at a temperature of 1250° C. for 70 minutes, and then cast into the crucible and cooled to obtain smelting slag and ruthenium alloy. The ruthenium content of the smelting slag is 4.3g; after calculation, the addition of the slag-forming agent meets the following conditions:

[0109] Calcium-boron ratio CaO: B2O3 = 2.69: 1, boron-silicon ratio B2O3: SiO2 = 0.23: 1, calcium-silicon ratio CaO: SiO2 = 0.62: 1, basicity = 0.71, melt theoretical viscosity = 6.29, melt theoretical surface tension = 0.34;

[0110] Step 3: The cobalt alloy and the aluminum block are configured in a mass ratio of 1:3, activated at a temperature of 1200°C for 60 minutes to obtain thermite alloy, and a 10 mol / L acid solution is added to the thermite alloy at a solid-liquid ratio of 1g:10ml to perform acid dissolution and impurity removal at a temperature of 95°C±2°C, and then solid-liquid separation is performed to obtain an acid leaching solution and an acid leaching residue;

[0111] Step 4: Put the acid leaching residue into a tubular furnace, introduce hydrogen at a gas flow rate of 4 L / min, and heat to 600°C at a heating rate of 6°C / min and keep warm for 5 hours to obtain ruthenium powder. After testing, the ruthenium content in the ruthenium powder is 1.0 g.

[0112] The ruthenium recovery amounts finally obtained in Examples 1-4 and Comparative Examples 1-4 were statistically analyzed and the ruthenium recovery rates were calculated. The results are shown in Table 1. The smelting slag and ruthenium alloy obtained in Example 1 and Comparative Examples 3 and 4 were photographed. Figure 1-4 ,in Figure 1 is a photo of the copper alloy of Example 1, Figure 2 This is a photo of the smelting slag of Comparative Example 3. Figure 3 This is a photo of the zirconium alloy of Comparative Example 3. Figure 4 This is a photo of the zirconium alloy of Comparative Example 4.

[0113] Table 1 Results of ruthenium loss, ruthenium recovery, and ruthenium recovery rate of Examples 1-4 and Comparative Examples 1-4;

[0114]

[0115]

[0116] According to Table 1 and Figure 1-4 The results show that:

[0117] According to the data comparison between Examples 1-4 and Comparative Examples 1-4, it can be seen that the present invention optimizes the calcium-boron ratio in the slag-forming agent on the basis of capturing and recovering ruthenium by the matte smelting method, thereby improving the capture effect of ruthenium and improving the recovery rate of ruthenium; wherein according to the data comparison between Examples 1-3 and Example 4, it can be seen that the present invention further optimizes the dosage of each component of the slag-forming agent on the basis of optimizing the calcium-boron ratio in the slag-forming agent, thereby controlling the boron-silicon ratio, calcium-silicon ratio, alkalinity, viscosity and surface tension of the melt, and the combination of the five produces a synergistic effect of improving the capture efficiency of ruthenium, and synergistically controls the formation of the slag phase and the coercion alloy phase, effectively improving the separation effect of the slag phase and the coercion alloy, and further improving the recovery rate of ruthenium.

[0118] According to the data comparison between Examples 1-4 and Comparative Example 1, it can be seen that the calcium-boron ratio produced by the slag-forming agent combination used in Comparative Example 1 is too large. Although the boron-silicon ratio, calcium-silicon ratio, alkalinity, and theoretical surface tension of the melt are all within the scope of the further optimized solution of the present invention, the excessive calcium-boron ratio affects the growth efficiency of calcium borate and the formation of the final slag phase, thereby affecting the separation effect of the slag phase and the ruthenium alloy, resulting in part of the ruthenium entering the slag phase during the capture process of free ruthenium ions, greatly reducing the recovery rate of ruthenium.

[0119] According to the data comparison between Examples 1-4 and Comparative Example 2, it can be seen that the calcium-boron ratio produced by the slag-forming agent combination used in Comparative Example 2 is too small. Although the boron-silicon ratio, calcium-silicon ratio, alkalinity, and theoretical surface tension of the melt are all within the scope of the further optimized solution of the present invention, the too small calcium-boron ratio affects the activation efficiency of calcium, thereby affecting the size of the internal driving force received by the ruthenium ions during the freeing process, and ultimately affecting the recovery rate of ruthenium; in this process, an inappropriate calcium-boron ratio may also affect the separation effect of the slag phase and the thorium alloy, further affecting the recovery rate of ruthenium. Therefore, the ruthenium recovery rate obtained in Comparative Example 2 is worse than that of the technical solution with an excessively large calcium-boron ratio in the comparative example.

[0120] According to the comparison of the data and pictures of Example 1 and Comparative Example 3, it can be seen that the technical solution of Comparative Example 3 adopts too large a calcium-boron ratio, too small a boron-silicon ratio, alkalinity, and too large a theoretical viscosity of the melt, which will cause the melt to have very poor fluidity, resulting in incomplete separation of the smelting slag and the cobalt alloy, a large amount of ruthenium remaining in the smelting slag, and other boron elements, calcium elements, and silicon elements in the cobalt alloy that wrap the ruthenium, making it difficult to carry out the subsequent ruthenium recovery treatment of the cobalt alloy, and the ruthenium recovery rate is greatly reduced; it can be seen that Figure 2 , a large amount of smelting slag hangs on the wall. If the smelting slag and the cobalt alloy are completely separated, the cobalt alloy will be at the bottom, and Figure 2 A large amount of smelting slag hanging on the wall proves that there is a problem with the formation of the cobalt alloy and it cannot be completely at the bottom. The separation effect of smelting slag and cobalt alloy is poor; and according to Figure 1 and Figure 3 From the comparison of the pictures of the zirconium alloys, it can be seen that the zirconium alloy obtained in Comparative Example 3 is irregular, indicating that the zirconium alloy obtained has too many impurities.

[0121] According to the data and pictures of Example 1, Comparative Example 3 and Comparative Example 4, when the theoretical viscosity of the melt is too large and exceeds a certain value, the recovery rate of ruthenium will be further reduced. Figure 4 It can be seen that the obtained zirconium alloy is not only irregular, but also has small droplets of alloy, which means that the separation effect of the zirconium alloy and the smelting slag is further deteriorated.

[0122] It should be noted that some ruthenium ions are still in the acid leaching solution of step 3 and need to be further extracted.

[0123] The embodiments presented herein are only embodiments selected according to the combination of all possible embodiments. The appended claims should not be limited by the embodiments describing the present invention. Some numerical ranges used in the claims include sub-ranges therein, and the changes in these ranges should also be covered by the appended claims.

Claims

1. A method for capturing and recovering ruthenium from ruthenium-containing waste, characterized in that: The steps include: Step 1: acid dissolving and removing impurities from the ruthenium-containing waste material in sequence, and separating to obtain a ruthenium-containing filter residue; Step 2: adding a collector, a slag-making agent and a reducing agent to the ruthenium-containing filter residue, and then performing matte smelting to obtain smelting slag and a cobalt alloy; Step 3: subjecting the cobalt alloy to activation treatment, acid dissolution and impurity removal, and calcination in sequence to obtain ruthenium powder; Wherein, the slag-forming agent includes calcium oxide and / or calcium carbonate, silicon dioxide, borax, and soda ash, and the usage ratio of calcium element to boron element in the slag-forming agent is CaO:B2O3=1-3:1 by mass.

2. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 1, characterized in that: The amount of calcium oxide and / or calcium carbonate, silicon dioxide, borax and soda ash in the slagging agent must meet the following conditions: The usage ratio of boron element to silicon element is B2O3:SiO2=0.2-0.6:1 by mass, the usage ratio of calcium element to silicon element is CaO:SiO2=0.5-1.2:1 by mass, the basicity is 0.9-1.5, the theoretical viscosity of the melt is 0.5-1.1, and the theoretical surface tension of the melt is 0.1-0.

5.

3. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 1, characterized in that: The ruthenium-containing waste in step 1 is activated ruthenium-containing waste, and the activation step of the ruthenium-containing waste is: the ruthenium-containing waste and the aluminum block are configured in a mass ratio of 1:2-4, and activated at a temperature of 1000-1400°C for 60-90 minutes.

4. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 1, characterized in that: The specific operation of step 1 is: adding 6-12 mol / L acid solution to the ruthenium-containing waste material at a solid-liquid ratio of 1g:10-12ml to carry out acid dissolution and impurity removal at a temperature of 90-100°C, and then performing solid-liquid separation to obtain a ruthenium-containing filter residue.

5. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 1, characterized in that: The specific operation of step 2 is: stir and mix the ruthenium-containing filter residue with the slag-forming agent and the reducing agent and put them into a crucible, then add the collector and spread it on the crucible, put the crucible into a smelting furnace and smelt at a temperature of 1200-1400°C for 60-90 minutes, then cast and cool to obtain smelting slag and ruthenium alloy; wherein the amount of the reducing agent added is 20-30wt% of the ruthenium-containing filter residue, and the amount of the collector added is 30-50wt% of the ruthenium-containing filter residue.

6. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 1, characterized in that: The specific operation of the activation treatment in step 3 is: the cobalt alloy and the aluminum block are configured in a mass ratio of 1:2-4, and activated at a temperature of 1000-1400° C. for 60-90 minutes to obtain the thermite alloy.

7. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 6, characterized in that: The specific operation of acid dissolution and impurity removal in step 3 is: adding 6-12 mol / L acid solution to the aluminothermic alloy according to the solid-liquid ratio of 1g:10-12ml to carry out acid dissolution and impurity removal at a temperature of 90-100°C, and then separating the solid and liquid to obtain acid leaching solution and acid leaching residue.

8. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 7, characterized in that: The step 3 also includes reduction precipitation, and the specific operation of reduction precipitation is: adding hydrochloric acid or sodium hydroxide to the acid leaching solution to adjust the pH of the acid leaching solution to 4-5, adding a sulfide solution with a sulfide content of 20wt% to adjust the pH to 7-9, stirring at a constant temperature of 40-80°C for 50-60min, and separating the solid and liquid to obtain ruthenium-containing slag.

9. The method for capturing and recovering ruthenium from ruthenium-containing waste according to claim 7, characterized in that: The specific operation of calcination in step 3 is: putting the acid leaching residue into a tubular furnace, introducing hydrogen at a gas flow rate of 3-5L / min, and heating to 400-600°C at a heating rate of 5-10°C / min and keeping the temperature for 4-6h to obtain ruthenium powder.

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