JP2009024263A - Platinum group element recovery equipment - Google Patents

Abstract

In a dry recovery method of a platinum group element in which a platinum group element such as a waste catalyst is heated and melted together with a copper source material to absorb the platinum group element in a molten metal, the operability of the furnace Further improve the recovery rate of platinum group elements.
A material to be treated containing a platinum group element and a copper source material containing copper oxide are charged together with a flux component and a reducing agent into a closed electric furnace and melted, and an oxide-based molten slag is obtained. In the platinum group element recovery method, in which molten metal mainly composed of copper metal is allowed to settle below the layer and the platinum group element is concentrated in the molten metal that has settled downward, the copper content is reduced to less than 3.0% by weight. A platinum group element dry recovery method is characterized in that slag is discharged from the electric furnace and a granular copper source material having a diameter of 0.1 to 10 mm is used as the copper source material.
[Selection] Figure 1

Description

  The present invention relates to a method for recovering platinum group elements from various substances containing platinum group elements, such as used petrochemical catalysts, used automobile exhaust gas purification catalysts, used electronic boards and lead frames, etc. .

  Conventionally, as a method for recovering platinum group elements from used automobile exhaust gas purification catalysts (ceramic carrier catalysts and metal carrier catalysts of exhaust gas converters: these are called "waste catalyst for automobiles") There are a method of extracting a platinum group element with a solution in which an oxidizing agent is added to an acid, and a method of dissolving a carrier using sulfuric acid or the like and separating it from an undissolved platinum group element. The extraction rate was poor, and a large amount of acid was used to dissolve the carrier, which had problems in recovery rate and cost.

On the other hand, the recovery methods described in Japanese Patent Application Laid-Open Nos. 4-317423 and 2000-248322 by the present applicants use a platinum group element-containing material such as a waste catalyst for automobiles in a furnace as a copper source material. A characteristic dry process of transferring platinum group elements into molten metal (molten copper metal) by melting with (copper oxide and / or metallic copper) was performed. By combining the concentration process of phase-separating molten metal containing platinum group elements into molten oxide and molten metal in which platinum group elements are further concentrated by oxidizing the platinum group elements at a high yield and low cost. Can be recovered, and has an advantage over the wet method as an economical resource recovery method.
JP-A-4-317423 JP 2000-248322 A JP 61-168719 A JP-A-9-170743 JP 2003-89824 A

  The dry recovery method for transferring the platinum group element into the molten metal is very excellent in terms of high recovery rate and low cost, but the platinum group element is sufficiently melted in the operation of the melting process. It took some settling time to move to metal. That is, when a platinum group element-containing substance such as an automobile waste catalyst and a copper source material are put into an electric furnace in a solid state, the platinum group element is transferred to the molten metal while they are melted down. When the phase separation of metal and metal occurs, it is necessary to have a timing at which the platinum group element can move to the metal side. For this reason, it was necessary to provide a relatively long settling time (stationary time) for safety reasons. In addition, the furnace condition may change with each material input, and this may cause the timing for the platinum group element to fully migrate into the molten metal.

  For this reason, in order to efficiently transfer platinum group elements into molten metal, it was necessary to analyze the melting behavior and take appropriate measures. An object of the present invention is to satisfy such a demand, and even if the settling time is shortened in the dry recovery method, the platinum group element can be efficiently and stably transferred to the molten metal side. The purpose is to improve.

  According to the present invention made to achieve the above object, a material to be treated containing a platinum group element and a copper source material containing copper oxide are charged into a closed electric furnace together with a flux component and a reducing agent. In the platinum group element recovery method, the molten metal mainly composed of copper metal is allowed to settle below the molten slag layer mainly composed of oxide, and the platinum group element is concentrated in the molten metal deposited below. A method for recovering a platinum group element is provided, wherein molten slag having reduced to 3.0% by weight or less is discharged from the electric furnace. In this method, the copper source material charged into the electric furnace is preferably a granular material having an average particle size of 0.1 mm or more and 10 mm or less. It is preferable to maintain the pressure lower than the atmospheric pressure.

  Further, according to the present invention, a material to be treated containing a platinum group element and a copper source material containing copper oxide are charged together with a flux component and a reducing agent into a closed electric furnace and melted. The molten metal mainly composed of copper metal is allowed to settle below the molten slag layer, the platinum group element is concentrated in the molten metal that has settled below, and the molten metal enriched with the platinum group element is separated from the molten slag to obtain another metal. In a separate furnace, the molten metal is transferred in a molten state, and the molten metal is oxidized to separate the oxide-based slag layer into a molten metal layer in which the platinum group element is further concentrated. In this method, the molten slag layer whose copper content has been reduced to 3.0% by weight or less is discharged from the electric furnace, and the molten slag produced in the separate furnace is cooled by water from a high temperature state to reduce the diameter to 0. .1m It provides a method for recovering platinum group elements, characterized in that to obtain a copper source material containing the copper oxide having the above 10mm or less of the particle.

  Further, as an apparatus suitable for carrying out this platinum group element recovery method, according to the present invention, a furnace body having an internal volume substantially blocked from the outside air and an upper body portion of the furnace body are provided. A material input port and an exhaust port, at least two fluid discharge ports with different height levels provided in the lower body portion of the furnace body, a material input chute connected to the material input port, and an exhaust port It consists of a connected exhaust system and an electrode for energizing and heating the material charged in the furnace. The charging material consists of an oxide-based material accompanied by a platinum group element, copper oxide, a solid reducing material and a flux. A metal system with a high platinum group element concentration in a reducing atmosphere that is substantially cut off from outside air and melted in the furnace while exhausting the gas generated in the furnace with the exhaust device, and from the lower fluid outlet Fluid is removed from the higher fluid outlet. Providing recovery apparatus platinum group element to extract the elemental concentrations of low slag based fluids.

  According to the present invention, the dry process of concentrating platinum group elements into molten metal copper from the platinum group element-containing treated materials such as waste catalysts for automobile exhaust gas purification provides a high yield of platinum group elements while streamlining furnace operation. Therefore, platinum group elements can be recovered economically from waste resources.

  In the present invention, the platinum group element-containing material to be treated is, for example, a used petrochemical waste catalyst containing platinum, palladium, etc., a spent automobile exhaust gas purification waste catalyst containing platinum, palladium, rhodium, etc. Of course, it includes lot-out products and scraps obtained from the manufacturing process of these catalysts, and also includes used electronic boards, dental parts, lead frames and the like containing palladium. Such a material to be treated containing a platinum group element is usually in a state where a trace amount of the platinum group element is supported on a metal oxide or ceramics.

These platinum group element-containing substances to be treated are introduced into an electric furnace and melted together with a copper oxide-containing copper source material, a flux and a carbonaceous reducing agent, and a metal is formed under the formed oxide-based molten slag layer. The basic structure of the method of the present invention is to settle the molten metal layer mainly composed of copper and concentrate the platinum group element in the molten metal layer that has settled downward. In this case, in the present invention,
1. Use a closed electric furnace as the electric furnace,
2. Discharge molten slag from the electric furnace whose copper content is reduced to 3.0 wt% or less, preferably 2.0 wt%,
3. Use a granular material having a particle size of 0.1 mm or more and 10 mm or less as a copper source material containing copper oxide to be charged into an electric furnace.
4. Maintain the electric furnace pressure below atmospheric pressure from melting of the charge to discharging the molten slag.
Adopting a characteristic method.
Hereinafter, these matters specified by the present invention will be described.

  FIG. 1 shows an example of equipment for carrying out the method of the present invention. In FIG. 1, reference numeral 1 denotes a closed electric furnace that occupies the main part of the equipment of the present invention. The closed electric furnace 1 includes a furnace body 3 having a furnace internal volume 2 substantially blocked from outside air, a material input port 4 and an exhaust port 5 provided in an upper body portion of the furnace body 3, and the furnace At least two fluid discharge ports 6 and 7 having different height levels provided in the lower body portion of the body 3, a material input chute 8 connected to the material input port 4, and an exhaust gas connected to the exhaust port 5 It comprises an apparatus 9 and electrodes 10a, 10b, 10c for energizing and heating the material charged in the furnace.

  The furnace body 3 shown in the figure is a furnace having a circular inner wall lined with a refractory, and a material charging chute 8 is arranged almost at the center of the ceiling surface, on a concentric circle centering on the material charging chute 8. The three electrodes 10a, 10b, and 10c are arranged from the ceiling surface at regular intervals and vertically. In other words, the vertical electrodes 10a, 10b, and 10c are arranged at the vertices of the equilateral triangle, and the material charging chute 8 exists at the approximate center of the equilateral triangle.

  The sealed electric furnace 1 configured in this way includes a granular material 11 containing a platinum group element, a granular copper source material 12 containing copper oxide, a solid reducing material (powder coke) 13 and a powder. The flux 14 is mixed and charged. That is, these charged raw materials are weighed and cut out from the respective hoppers and fed into the material charging chute 8 of the electric furnace 1 while being mixed and conveyed by the screw conveyor 15. The material charging chute 8 is provided with two upper and lower shutters 16 and 17 so that airtightness is maintained when the material is charged into the furnace. First, the upper shutter 16 is opened, the lower shutter 17 is closed, and one batch of material is charged into the chute 8, then the upper shutter 16 is closed, the lower shutter 17 is opened, and stored in the chute 8. The batch that has been put into the furnace. When the batch is charged into the furnace, the upper and lower shutters 16 and 17 are closed to prepare for the next charging operation. In the facility shown in the figure, a branch member 18 is attached to the lower end of the chute 8 (below the material inlet 4), so that the material falling from the chute 8 into the furnace connects the three electrodes 10a, 10b, 10c. In the vicinity of each side, more preferably, near the midpoint position of each of the three sides. As a result, new input materials are deposited at the shortest distance connecting the electrodes 10a, 10b, and 10c, and the melting efficiency is increased.

  At the initial stage of operation of the furnace in which new material is introduced, at least two fluid discharge ports 6 and 7 provided at the lower body portion of the furnace body 3 and having different height levels are closed. When the electrodes 10a, 10b, and 10c are energized, the charged substance in the furnace starts to melt. During this time, the exhaust gas generated in the furnace by driving the exhaust device 9 is discharged from the exhaust port 5 through the dust removing device 19. Is exhausted outside the system after the exhaust gas treatment. By the continuous operation of the exhaust device 9, the furnace space 2 that is substantially shut off from the outside air is maintained at a pressure lower than the atmospheric pressure.

  When the material charged in the furnace begins to melt, the metal oxide, especially the copper oxide in the copper source material, is reduced to metal copper by the reducing material (powder coke) to form molten metal copper. Since the metal melt has a higher specific gravity than the oxide melt (slag), the metal melt descends in the slag and sinks below the furnace to form a molten metal pool 20. On the molten metal 20, an oxide melt, that is, a slag layer 21 is formed.

  In the process in which the molten metal copper produced by reduction of copper oxide descends in the slag, the platinum group elements present in the slag are taken into the molten metal copper. That is, let it melt. As a result, the molten metal 20 is collected in a state where the platinum group element is dissolved in the molten metal 20, and the molten metal 20 having a high platinum group element concentration is obtained. On the other hand, the concentration of the platinum group element in the slag 21 is decreased by the amount of the platinum group element dissolved in the molten metal 20. Accordingly, the slag 21 having a lower platinum group element concentration is separated from the higher fluid discharge port 6 and the molten metal 20 having a higher platinum group element concentration is separated from the lower fluid discharge port 7 while being separated from each other. When it is allowed to flow out, it is possible to collect a molten metal having a high platinum group element concentration (metal in which a platinum group element is dissolved in metallic copper).

  In this way, by performing the platinum group element recovery process using the closed electric furnace 1, the molten metal of the platinum group element is maintained in a reducing atmosphere and with high thermal efficiency. As a result, the processing time can be shortened and the platinum group element recovery rate can be improved.

  One feature of the method of the present invention is that, when such a closed electric furnace is used to recover platinum group elements, molten slag having a copper content reduced to 3.0% by weight or less is removed from the electric furnace. It is in point to discharge. As shown in the examples described later, about 0.3 ton of copper oxide is used for 1 ton of a platinum group element-containing material, and the raw material is mixed with a reducing material sufficient to reduce all of the copper oxide. When blending (mixing nearly 1 ton of flux component), the platinum group element content remaining in the slag is the content of copper remaining in the slag when smelting reduction is performed in a closed electric furnace. It was found to be closely related to the amount. The relationship is shown in FIG.

  As shown in FIG. 2, when the Cu content in the slag is, for example, 1% by weight, the Pt, Pd and Rh in the slag are reduced to about 5 ppm, about 3 ppm and about 1 ppm, respectively. It can be seen that the contents of Pt, Pd, and Rh all tend to further decrease as the temperature decreases further. However, in the region where the Cu content exceeds 3.0% by weight, the contents of Pt, Pd and Rh all show a tendency to increase suddenly, and the recovery rate of the platinum group element decreases rapidly.

  Therefore, when discharging the molten slag from the electric furnace, it is desirable to discharge the molten slag whose copper content is reduced to 3.0% by weight or less, preferably 2.0% by weight or less. The content of copper in the molten slag can be known in real time by sampling the slag in the furnace during operation and analyzing it by equipment.

  In the practice of the present invention, when the amount of molten slag increases and it becomes necessary to discharge it outside the furnace, the copper content in the slag is measured and exceeds 3.0% by weight. In this case, it is preferable that the contents in the furnace are allowed to stand under a predetermined temperature condition without performing the discharging operation. During this standing, the copper content in the slag gradually decreases, and accordingly, the platinum group elements in the slag also move to the molten metal side.

  In this way, even if slag with a low copper content is discharged outside the furnace, platinum group elements can be prevented from flowing out of the furnace along with this, and as a result, the molten metal with a high platinum group element concentration can be avoided. Can be recovered.

  Another feature of the method of the present invention is that a granular material having a particle size of 0.1 mm or more and 10 mm or less is used as a copper oxide-containing copper source material charged in an electric furnace. When a granular material having a particle size of 0.1 mm or more and 10 mm or less is used as the copper source material, the platinum group element in the material to be treated easily migrates into the molten metal when the material to be treated and the copper source material are heated and melted. I found out that In particular, it is desirable that a copper source material having a particle size of 0.1 mm or more and 10 mm or less is present in an amount of 50% by weight or more. If the condition is satisfied, other materials may be a lump of 10 mm or more. In some cases, a powder of less than 0.1 mm may be mixed.

  It is preferable that at least 50% by weight or more of the platinum group element-containing substance to be treated is a granular material having a particle size of 10 mm or less in order to improve the mixing property with the copper source material. Both the material to be treated and the copper source material are granular bodies having appropriate particle sizes. When this material is mixed with the carbonaceous reductant and the flux and charged into the furnace, the copper oxide in the copper source material melts. -It becomes easy to be reduced, and the opportunity for the produced molten metal copper to come into contact with the platinum group element in the material to be treated in the vicinity increases, so that the platinum group element is incorporated into the molten metal copper.

In order to promote the meltdown of the material to be treated and the copper source material and to improve the fluidity of the slag to be produced, it is desirable to add the flux simultaneously to the charging material. As the flux, silica, calcium oxide, calcium carbonate or the like mixed at an appropriate ratio is preferable. The mixing ratio of the flux components varies depending on the composition of the raw material, but the composition of the slag after heating and melting is as follows: Al ₂ O ₃ : 20 to 40% by weight, SiO ₂ : 25 to 35% by weight, CaO: 20 to 30% by weight, FeO : It is preferable to mix | blend a flux component with a charging raw material so that it may become 5 to 30 weight%.

  In order to reduce the copper oxide in the copper source material to obtain a molten metal of metallic copper, coke is preferably added as a reducing agent. However, in addition to coke, a base metal containing a valuable metal having a reducing action and a carbon source Other resin materials, activated carbon, etc. can also be used. According to the present invention, valuable metals (noble metals and platinum group elements) contained in these reducing agents can also be recovered simultaneously.

  In carrying out the method of the present invention, a sealed electric furnace is charged with a mixture of a substance to be treated, a copper source material, a flux and a reducing agent, and the furnace pressure is maintained at a pressure slightly lower than atmospheric pressure. It is heated and melted at a temperature of 1100 ° C. to 1700 ° C., more preferably 1300 ° C. to 1500 ° C., the oxide in the charging material is melted, and the copper oxide in the charging material is reduced to copper. If the heating and melting temperature is less than 1100 ° C, the molten state of the slag is not perfect and the viscosity increases, and the recovery rate of the platinum group elements may be reduced. It becomes a factor inviting. By maintaining the inside of the furnace under reduced pressure, the reducing atmosphere is maintained, the reduction of copper oxide to copper proceeds well, and the efficiency of absorption of platinum group elements into molten metal is also increased.

  In the melt-down state of the charged material, most of the material to be treated becomes a glassy molten oxide layer (slag layer). Copper oxide is reduced by a reducing agent to become molten metal copper. Both are naturally separated into two layers due to the difference in specific gravity, and a slag layer is formed in the upper layer and a molten metal layer is formed in the lower layer. At this time, the platinum group element in the raw material to be treated moves to the lower molten metal layer and is absorbed, but as described above, the grain size of the copper source material is absorbed by the molten metal layer and shortening its settling time. This greatly affects the improvement of the yield of the platinum group elements, and when the particle size of the copper source material is set to 0.1 mm or more and less than 10 mm, a remarkable effect appears on these improvements.

  The reason is not necessarily clear, but can be considered as follows. The platinum group element in the material to be treated is dispersed in slag having an appropriate viscosity when the material to be treated is melted down together with the flux. Immediately after the copper oxide added at the same time is reduced, it is dispersed as molten metal in the slag, and the platinum group elements dispersed and suspended in the slag having an appropriate viscosity are absorbed in the slag layer. Descend. The inventors named the behavior of the molten metal (copper metal) absorbing the platinum group element as "copper showering effect". If the particle diameter of the initially supplied copper source material is less than 0.1 mm, the molten metal copper dispersed in the slag is also fine, so that it often settles down to the lower metal layer. It takes time and the copper showering effect does not work sufficiently. On the other hand, if the diameter of the initially introduced copper source material exceeds 10 mm, the molten metal copper is absorbed into the lower metal layer before sufficiently absorbing the platinum group elements dispersed in the slag. In this case as well, the copper showering effect does not function sufficiently. In addition, a certain amount of surface area and cross-sectional area are required for the descending molten metal copper to absorb the platinum group elements dispersed in the slag. That is, the absorption efficiency increases as the surface area and cross-sectional area increase even when the weight of the copper source material to be added is the same. For these reasons, the copper showering effect works most efficiently when the particle size of the initially introduced copper source material is 0.1 mm or more and 10 mm or less, and it melts from the melted down material to be treated. It is considered that the migration of the platinum group element into the metal is performed well.

  According to the experience of the inventors, if 50% by weight or more, preferably 80% by weight or more of the copper source material has a particle size in this range, there is substantially no problem in the recovery of the platinum group element. When the particle size was less than 50% by weight, it was necessary to take a long time for standing, that is, settling time, in order to increase the recovery rate of the platinum group element. Here, stationary, that is, set ring, means that the slag that has already melted after the material is charged into the electric furnace is energized as it is in order to maintain it at a predetermined temperature. Meanwhile, the pressure in the closed electric furnace is preferably maintained under reduced pressure.

  After this standing, most of the upper slag is discharged out of the furnace, leaving a part of it in the furnace when the copper content becomes 3.0% by weight or less as described above. The molten metal layer that has absorbed the platinum group elements existing in the lower layer in the furnace is also tapped outside the furnace while leaving a part of it in the furnace. Although the molten slag and other parts of the molten metal remain in the furnace, the same operation can be repeated again by charging the material of the next heat into the furnace in this state.

  The molten metal, which is extracted from the sealed electric furnace and separated from the molten slag, is concentrated. The molten metal is transferred to the oxidation furnace in the molten state, and the platinum group element is further concentrated in the molten metal. Good to do.

  In the oxidation furnace, this molten metal is oxidized in the molten state, and the molten oxide (copper oxide) formed on the molten metal surface is discharged outside the furnace, leaving the molten metal further enriched with platinum group elements. That is, the platinum group element hardly migrates to the molten oxide layer formed on the molten metal surface and remains in the lower molten metal layer. Therefore, every time the generated molten oxide layer is discharged, The platinum group element concentration increases. The oxidation treatment in this oxidation furnace is preferably carried out by introducing an oxygen gas or an oxygen-containing gas while maintaining the material temperature at a temperature of 1100 ° C. to 1700 ° C., preferably 1200 ° C. to 1500 ° C. If it is less than 1100 ° C., solidification of the molten oxide or molten metal occurs, and the progress of oxidation is inhibited. If the temperature exceeds 1700 ° C., the furnace body will be damaged.

  In this way, by repeating the oxidation process and the oxide layer discharge process in the oxidation furnace, the molten metal layer enriched with the platinum group element increases the platinum group element content to 10 to 75% by weight. Can do. After taking it out of the oxidation furnace, it is sent to the platinum group element recovery and purification in the next step to separate and refine metallic copper and platinum group elements.

  On the other hand, the molten oxide layer (oxide mainly composed of copper oxide) discharged from the oxidation furnace can be reused as a copper source material charged in the electric furnace. At that time, by putting the oxide layer discharged from the oxidation furnace in a molten state into water, that is, by granulating, the granular material having a particle size of 0.1 mm or more and 10 mm or less is preferably 50% by weight or more. A copper source material containing 80% by weight or more can be obtained. The obtained granulated water can be dried and then further sized with a sieve or the like to obtain a copper source material having a particle size suitable for the treatment of the present invention. This copper source material inevitably accompanies platinum group elements, but by reusing this, the accompanying platinum group elements will eventually migrate into the molten metal layer, which will further increase the recovery rate of platinum group elements. Become.

  The present invention will be further described below with reference to examples of the present invention.

[Example 1]
Waste catalyst for purification of automobile exhaust gas containing Pt: 1200 ppm, Pd: 450 ppm, Rh: 90 ppm (Al ₂ O ₃ : 36.5 wt%, SiO ₂ : 40.6 wt%, MgO: 10.5) Containing 10% by weight) was crushed to 10 mm or less. Further, copper oxide containing 80% by weight of a granular material having a particle size of 0.1 mm or more and 10 mm or less (the remainder is a massive copper oxide having a particle size exceeding 10 mm) was prepared as a copper source material. 300 kg of the copper source material containing this granular material was mixed with 1000 kg of the substance to be treated, and further 600 kg of CaO, 200 kg of Fe ₂ O ₃ and 400 kg of SiO ₂ were mixed as flux components, and 30 kg of coke as a reducing agent.

  This mixture was put into a closed electric furnace as shown in FIG. 1 and heated and melted at 1350 ° C. In the electric furnace at the time of charging the mixture, the previously melted molten metal and the molten slag remain on the upper part, and the molten slag is the remaining 1 after the about 3/4 of the previous melt was discharged. / 4 remains.

  After charging the mixture, the charge is heated and melted at 1350 ° C. while driving the exhaust system and maintaining the inside of the furnace at a reduced pressure. When the charged mixture floating on the slag surface melts, the slag is sampled. When the copper content was analyzed, it was 0.8% by weight. For this reason, about 3/4 of the slag layer was immediately discharged from the side of the electric furnace. When the amount of the platinum group element in the slag that was removed and cooled and solidified was analyzed, it was found that Pt: 0.7 ppm, Pd: 0.1 ppm, Rh: 0.1 ppm or less. That is, most of the platinum group elements moved to the molten metal layer under the electric furnace.

[Example 2]
Example 1 was repeated except that copper oxide containing 50% by weight of particles having a particle size of 0.1 mm to 10 mm was used as the copper source material (the remainder was a massive copper oxide having a particle size exceeding 10 mm). It was. As a result, the copper content at the time of slag discharge was 0.9% by weight, and the platinum group elements in the slag were Pt: 0.9 ppm, Pd: 0.2 ppm, and Rh: 0.1 ppm or less.

[Comparative Example 1]
Example 1 was repeated except that 15 kg of coke powder as the reducing material was used. As a result, the content of copper in the slag at the time of slag discharge was 3.2% by weight, and the platinum group elements in the slag were Pt: 20 ppm, Pd: 12 ppm, and Rh: 2 ppm.

[Comparative Example 2]
Example 1 was repeated, except that copper oxide containing 60% by weight of powder having a particle size of less than 0.1 mm (the remaining copper oxide having a particle size of 0.1 mm or more) was used as the copper source material. As a result, platinum group elements in the slag were Pt: 3.8 ppm, Pd: 1.2 ppm, and Rh: 0.2 ppm.

[Comparative Example 3]
Example 1 was used except that the copper source material contained 30% by weight of a granular material having a particle size of 0.1 mm or more and 10 mm or less, and the remaining 70% by weight was a copper oxide having a lump with a diameter exceeding 10 mm. Repeated. As a result, platinum group elements in the slag were Pt: 4.2 ppm, Pd: 1.6 ppm, and Rh: 0.2 ppm.

Example 3
After the discharge in Example 1, about 2/3 of the molten metal was discharged from the lower part of the electric furnace, and the molten metal was charged into the oxidation furnace in a molten state. In this oxidation furnace, oxygen-enriched air having an oxygen concentration of 40% was blown from the top blowing lance onto the surface of the molten metal. When the oxide layer was formed to a thickness of about 1 cm on the surface of the molten metal, the furnace was tilted to cause the oxide (copper oxide) layer to flow out of the furnace and put into a water tank in which a large amount of water flowed.

  Subsequently, oxygen-enriched air was blown onto the molten metal layer in the oxidation furnace, and when the oxide layer was formed to about 1 cm, the furnace was tilted to discharge the oxide in the same manner, and the operation was repeated. . Thereafter, the water-crushed oxide (copper oxide-based substance) was taken out of the water tank, dried, sampled, and the particle size and composition were measured with a sieve. As a result, the particulate matter having a particle size of 0.1 mm or more and 10 mm or less was 99% by weight.

Example 4
After the copper oxide was discharged from the oxidation furnace in Example 3, the molten metal present in the lower part of the electric furnace after discharging of Example 2 was discharged on the molten metal layer remaining in the oxidation furnace. I was charged. Then, oxidation treatment was performed in the same manner as in Example 3 to obtain a water-crushed oxide. As a result, the particle size of the water-crushed oxide (a material mainly composed of copper oxide) was 0.1 mm or more and 10 mm or less. The granular material was 99% by weight.

  The entire molten metal layer present in the lower layer of the oxidation furnace was taken out and solidified by cooling, and 10 kg of metallic copper enriched with platinum group elements was collected. The platinum group element content in the copper metal was Pt: 23 wt%, Pd: 8.5 wt%, and Rh: 1.5 wt%.

Example 5
Example 1 was repeated except that instead of the copper oxide of Example 1, the crushed oxide obtained in Example 3 (substance based on copper oxide) was used. The copper content at the time of slag discharge was 0.8% by weight, and the platinum group element content in the obtained slag was Pt: 0.7 ppm, Pd: 0.1 ppm, Rh: 0.1 ppm or less. It was.

It is a schematic sectional drawing which shows the example of the apparatus which enforces this invention method. It is a figure which shows the relationship between the copper content in slag at the time of implementing this invention, and platinum group element content.

Explanation of symbols

DESCRIPTION OF SYMBOLS 1 Sealed electric furnace 2 Furnace volume 3 Furnace body 4 Material inlet 5 Exhaust port 6 Higher fluid outlet 7 Lower fluid outlet 8 Material input chute 9 Exhaust device 10 Electrode

Claims (3)

  A furnace body having an internal volume substantially blocked from outside air, a material input port and an exhaust port provided in the upper body part of the furnace body, and at least different height levels provided in the lower body part of the furnace body It consists of two fluid discharge ports, a material input chute connected to the material input port, an exhaust device connected to the exhaust port, and an electrode for energizing and heating the material charged in the furnace. , A platinum group element-containing oxide-based material, copper oxide, a solid reducing material, and a charging material in a reducing atmosphere, which is substantially shielded from the outside air, and exhausting the gas generated in the furnace with the exhaust device It melts in the furnace, and a metal-based fluid with a high platinum group element concentration is taken out from the lower fluid outlet, and a slag fluid with a low platinum group element concentration is taken out from the higher fluid outlet. The platinum group element recovery device.   The platinum group element recovery device according to claim 1, wherein a mixture of the oxide-based raw material accompanied by the granular platinum group element, the granular copper oxide, the powdered solid reducing material, and the powdered flux is led to the material charging chute.   Remove only part of the metal fluid with high platinum group element concentration from the lower fluid outlet and remove only part of the slag fluid with lower platinum group element concentration from the higher fluid outlet. 2. The platinum group element recovery device according to claim 1, wherein after taking out, a new charging material is charged from a material charging chute while other fluid remains in the furnace, and melting is continued.

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