JP7400443B2 - Mutual separation method of platinum group elements - Google Patents

Description

The present invention relates to a method for mutually separating platinum group elements, and more particularly to a method that can further improve the recovery rate of platinum (Pt).

Platinum group elements are rare elements in terms of resources, and are rarely produced in natural minerals such as platinum ore, which contain high concentrations of platinum group elements.The raw materials for industrially produced platinum group elements are copper, The bulk of this comes from by-products from the smelting of non-ferrous metals such as nickel and cobalt, as well as various used waste catalysts such as automobile exhaust gas treatment catalysts.

By-products from non-ferrous metal smelting are platinum group elements such as platinum, palladium, iridium, rhodium, ruthenium, and osmium, which are contained in very small amounts in the smelting raw materials, and are the main metals due to their chemical properties. It is concentrated in sulfide concentrates and crude metals such as copper and nickel, and is further separated in the form of precious metal concentrates containing platinum group elements as residues in the main metal recovery process such as electrolytic refining.

In this concentrate, along with the main metals such as copper and nickel, other constituent elements such as precious metals such as gold and silver, group VI elements such as selenium and tellurium, and group V elements such as arsenic are combined into platinum group elements. Generally, they coexist in a relatively high content. Thereafter, gold and silver are recovered to obtain a platinum group element-containing material containing impurity elements, and the platinum group elements are separated and recovered. In a method for industrially separating platinum group elements from platinum group element-containing materials, the platinum group elements are usually leached into a liquid and then mutually separated and purified using separation techniques such as solvent extraction and adsorbent, and then recovered.

Here, Patent Document 1 describes a method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements, using a highly safe compound and process to remove impurities from platinum group elements in a mother liquor. It prevents the increase in the content ratio of elements and the decomposition of chloro complexes, effectively removes impurity elements, and removes palladium (Pd), platinum (Pt), iridium (Ir), ruthenium (Ru), and rhodium (Rh). ) are shown to be mutually separated at a purity that can be commercialized. Specifically, by processing PGMconc (noble metals are concentrated by processing anode slime) according to the flow shown in FIG. 4, it is possible to mutually separate platinum group elements at a purity that can be commercialized.

However, in the method disclosed in Patent Document 1, when Pt in the Bi-SX raffinate shown in FIG. 4 is 100, the oxidized Pt in the solution after neutralization is about 65. Therefore, in order to further improve the Pt recovery rate, in the final stage of the flow, the Ru residue solution is oxidized and neutralized, and the resulting oxidized and neutralized solution is returned to the Pt purification process (see Figure 4). Flow arrow X) is required.

However, as for Pt contained in the IrRuRh precipitate, loss occurs during the Ru leaching process, hydrochloric acid dissolution process, and oxidation neutralization process, and even if the route for returning it as described above is provided, the Pt purification process The amount of Pt introduced into the sample remained at less than 90 (when Pt in the Bi-SX raffinate was taken as 100).

Japanese Patent Application Publication No. 2005-097695

The present invention has been made in view of the above circumstances, and is capable of improving the recovery rate of platinum (Pt) in a method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements. The purpose is to provide a method.

The present inventors have made extensive studies to solve the above-mentioned problems. As a result, the leachate obtained by leaching the platinum group element containing impurity elements with hydrochloric acid is subjected to solvent extraction treatment, and potassium chloride is added at a predetermined ratio to the extraction residue to form a precipitate. The inventors have discovered that by doing so, Pt can be effectively transferred into the precipitate and the recovery rate of Pt can be improved, and the present invention has been completed.

(1) The first invention of the present invention is a method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements, wherein the platinum group element-containing material is suspended in a hydrochloric acid solution and an oxidizing agent is added. a leaching step to obtain a leached product solution containing a platinum group element, a solvent extraction step in which the leached product solution is subjected to a solvent extraction process, and an extraction residue obtained by the solvent extraction process, This is a method for mutually separating platinum group elements, including a caking step in which potassium chloride is added to the extraction residue at a concentration of 20 g/L or more to form a precipitate.

(2) A second invention of the present invention is the mutual interaction of platinum group elements in the first invention, wherein in the caking step, the potassium chloride is added to the extraction residue in an amount of 80 g/L or less. This is a separation method.

(3) In the third invention of the present invention, in the first or second invention, the solvent extraction step includes contacting the leaching product solution with diethylene glycol dibutyl ether and subjecting it to solvent extraction, and extracting organic matter containing impurity elements. The first step is to form a phase and an extraction residue, and the resulting extraction residue is brought into contact with an alkyl sulfide and subjected to solvent extraction to extract palladium, and then back-extracted to form a back-extraction product solution containing palladium. a second step of forming an extraction residue; the extraction residue obtained in the second step is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction; This method includes a third step of forming a phase and an extraction residue, and the extraction residue obtained in the third step is subjected to the caking step.

(4) In the fourth invention of the present invention, in the first or second invention, the solvent extraction step includes contacting the leaching product solution with diethylene glycol dibutyl ether and subjecting it to solvent extraction, and extracting organic matter containing impurity elements. The first step is to form a phase and an extraction residue, and the resulting extraction residue is brought into contact with an alkyl sulfide and subjected to solvent extraction to extract palladium, and then back-extracted to form a back-extraction product solution containing palladium. and a second step of forming an extraction residue, and the extraction residue obtained in the second step is subjected to the caking step.

(5) A fifth invention of the present invention is a platinum group metal refining process according to any one of the first to fourth inventions, further comprising a platinum refining process of refining platinum contained in the precipitate obtained in the caking process. This is a method of mutually separating elements.

(6) A sixth invention of the present invention is the method according to any one of the first to fifth inventions, further comprising adjusting the pH of the filtrate produced by filtering and separating the precipitate obtained in the caking step to 5 to 12. An oxidation-neutralization step in which the precipitates containing iridium, ruthenium, and rhodium are produced by adding an oxidizing agent and subjecting them to hydrolysis; a ruthenium leaching step in which an oxidizing agent is added and leached to produce a residue containing iridium and rhodium and a ruthenium leaching product solution; and iridium and rhodium obtained by dissolving the residue obtained in the ruthenium leaching step in a hydrochloric acid solution. An iridium extraction process in which an aqueous solution containing rhodium is brought into contact with tributyl phosphate and subjected to solvent extraction to extract iridium, and then back extracted to produce a back extraction product solution containing iridium and an extraction residue containing rhodium. A method for mutually separating platinum group elements, comprising:

(7) In the seventh aspect of the present invention, in the fourth aspect, the pH of the filtrate produced by separating the precipitate obtained in the caking step by filtration is adjusted to 5 to 12, and the oxidizing agent is added to the filtrate. An oxidative neutralization step in which a precipitate containing iridium, ruthenium and rhodium is produced by adding and hydrolyzing the precipitate, and an oxidizing agent is added to the precipitate obtained in the oxidative neutralization step in an alkaline aqueous solution with a pH of 12 or higher. A ruthenium leaching step in which a ruthenium leaching process is performed to produce a residue containing iridium and rhodium and a ruthenium leaching product solution, and an aqueous solution containing iridium and rhodium obtained by dissolving the residue obtained in the ruthenium leaching process in a hydrochloric acid solution, an iridium extraction step of contacting with tributyl phosphate and subjecting it to solvent extraction to extract iridium, and then back-extracting it to produce a back-extraction product solution containing iridium and an extraction residual solution containing rhodium; and the iridium extraction step. Impurity removal involves bringing the resulting extraction residue into contact with bis(2-ethylhexyl) phosphoric acid and subjecting it to solvent extraction to form an organic phase containing cationic impurity elements and an extraction residue to remove impurities. A method for mutually separating platinum group elements, comprising the steps of:

According to the present invention, the recovery rate of platinum (Pt) can be improved in a method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements.

FIG. 2 is a process diagram showing an example of the flow of a method for mutually separating platinum group elements. FIG. 3 is a process diagram showing another example of the flow of the method for mutually separating platinum group elements. FIG. 3 is a process diagram showing another example of the flow of the method for mutually separating platinum group elements. 1 is a process diagram showing the flow of a conventional method for mutually separating platinum group elements.

Hereinafter, specific embodiments of the present invention will be described in detail. Note that the present invention is not limited to the following embodiments, and various changes can be made without changing the gist of the present invention. Furthermore, in this specification, when expressed as "X to Y" (X and Y are arbitrary numerical values), it means "more than or equal to X and less than or equal to Y" unless otherwise specified.

The method for mutually separating platinum group elements according to the present invention (hereinafter also simply referred to as "mutual separation method") is a method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements.

Platinum group element-containing materials containing impurity elements are not particularly limited, and include various impurities obtained from by-products from smelting of non-ferrous metals such as copper, nickel, and cobalt, various used waste catalysts such as automobile exhaust gas treatment catalysts, etc. Concentrates of platinum group elements containing the elements can be used. Examples of impurity elements include main metals such as copper, nickel, cobalt, and iron, as well as other constituent elements such as gold, silver, lead, tin, selenium, tellurium, arsenic, antimony, and bismuth. Examples of platinum group element-containing materials containing such impurity elements include those obtained by processing anode slime to concentrate noble metals (PGMconc).

Specifically, this mutual separation method includes a leaching step in which a platinum group element-containing material is leached with a hydrochloric acid solution to obtain a leaching product solution containing a platinum group element, and a solvent extraction step in which a solvent extraction process is performed on the leaching product solution. and a caking step of adding potassium chloride to the extraction residue obtained by the solvent extraction treatment to form a cake (precipitate).

According to such a method, most of the platinum (Pt) contained in the extraction residue obtained through the solvent extraction step (90% or more) can be transferred into the precipitate, and the recovery rate of Pt can be increased. can be improved efficiently and effectively. In addition, in conventional methods (for example, the method disclosed in Patent Document 1), Pt is concentrated in the post-treatment solution (post-oxidation-neutralization solution) obtained by subjecting the extraction residual solution to oxidation-neutralization treatment. A process such as returning the concentrated liquid to the platinum refining process is required, but the method of the present invention eliminates the need for such a process, making it possible to simplify the overall process. Efficiency can be increased.

FIG. 1 is a process diagram showing an example of the flow of a method for mutually separating platinum group elements. As described above, for example, as the platinum group element-containing substance containing impurity elements, anode slime treated to concentrate noble metals (PGMconc) can be used. Note that FIG. 2 is a process diagram showing another example of the flow of the method for mutually separating platinum group elements, and is different from the embodiment shown in FIG. 1 in terms of the processing flow in the solvent extraction step S2.

[Leaching process of platinum group element-containing substances]
In the mutual separation method according to the present invention, a platinum group element-containing material (PGMconc) containing impurity elements is subjected to a leaching step S1 using a hydrochloric acid solution. Specifically, in the leaching step S1, a platinum group element-containing material containing impurity elements is suspended in a hydrochloric acid solution, an oxidizing agent is added, and the suspension is subjected to leaching treatment to obtain a leaching product liquid containing a platinum group element.

Since the platinum group element-containing material to be treated usually contains platinum group elements in the form of metal or sulfide, it is treated by applying an oxidizing agent in the presence of a hydrochloric acid solution in the leaching step S1. Therefore, platinum group elements can be effectively dissolved.

Specifically, in the leaching step S1, a platinum group element-containing material is suspended in an aqueous solution containing hydrochloric acid, and an oxidizing agent is added thereto. Here, the hydrochloric acid solution may be made to coexist in the aqueous solution from the beginning, but if the raw material is a sulfide, it may be chemically generated by a reaction between chlorine and water. In the leaching process, most of the impurities associated with the platinum group elements, lead and silver, remain in the residue as chlorides, and all other elements are dissolved as chlorides or chloro complexes.

The oxidizing agent is not particularly limited, and includes, for example, nitric acid, hydrogen peroxide, chlorate, chlorite, hypochlorite, chlorine, bromate, hypobromite, bromine, peroxosulfate. etc. can be used. Among them, from a practical standpoint, at least one selected from nitric acid, hydrogen peroxide, and chlorine is preferable in terms of cost.

The conditions for the leaching treatment are not particularly limited, but are preferably conditions that can reliably convert the platinum group element into a chloro complex. For example, the temperature is preferably 70° C. or higher, and the hydrochloric acid concentration in the liquid portion of the suspension is preferably 4 mol/L or higher. In order to prevent hydrolysis of the platinum group element in the subsequent solvent extraction step S2 (particularly in the third step S23 described later), it is desirable to form a chloro complex reliably in the leaching treatment.

[Solvent extraction process]
In the mutual separation method according to the present invention, a solvent extraction step S2 (S2A, S2B) is performed in which a leaching product liquid obtained through the leaching step S1 is subjected to a solvent extraction process. This solvent extraction step S2 can be a continuous solvent extraction step in which a solvent extraction process described below is performed. Note that the solvent extraction process S2A is a first embodiment, and the solvent extraction process S2B is a second embodiment, and either solvent extraction process may be performed.

- First embodiment of the solvent extraction step Specifically, the solvent extraction step S2A includes the first step S21 in which the leaching product is brought into contact with diethylene glycol dibutyl ether and subjected to solvent extraction, and the extraction obtained from the first step S21. A second step S22 in which the residual liquid is brought into contact with an alkyl sulfide and subjected to solvent extraction, and a third step S22 in which the extracted residual liquid obtained from the second step S22 is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction. Step S23.

(1st step)
The first step S21 in the solvent extraction step S2A is a step of bringing the leaching product obtained in the leaching step S1 into contact with diethylene glycol dibutyl ether and subjecting it to solvent extraction to form an organic phase containing impurity elements and an extraction residue. . In this first step S21, among the impurity elements contained in the leaching solution, elements that form lipophilic chlorocomplexes such as gold, tin, antimony, tellurium, and iron are almost completely extracted, and trivalent arsenic and Tetravalent selenium can also be extracted. Thereby, impurity elements can be effectively aggregated and separated and removed.

Here, the concentration of hydrochloric acid in the leaching product liquid to be subjected to the solvent extraction treatment is not particularly limited, but it is preferably adjusted to about 4 mol/L to 9 mol/L. When the hydrochloric acid concentration is less than 4 mol/L, the extraction rate of impurities other than gold is greatly reduced. On the other hand, when the hydrochloric acid concentration exceeds 9 mol/L, the elution of diethylene glycol dibutyl ether into the aqueous phase increases significantly. A small portion of the platinum group elements is also extracted and enters the organic phase, but by washing the organic phase with an aqueous solution of hydrochloric acid in the concentration range mentioned above, the platinum group elements can be back-extracted and transferred into the aqueous phase. It can be recovered.

Recovery of the extracted impurity elements from the organic phase is not particularly limited and can be performed according to known methods. For example, gold can be isolated by back extraction with a reducing aqueous solution such as oxalic acid or sodium sulfite, and other impurity elements can be separated and removed from the organic phase as precipitates of hydroxides or basic salts. Furthermore, by maintaining the pH of the back extraction at -0.2 or lower, only gold can be recovered as a metal, and other impurities can be separated while being dissolved in the back extraction product liquid.

Such solvent extraction treatment in the first step S21 can prevent an increase in the content ratio of impurity elements to platinum group elements in the mother liquor.

(Second process)
In the second step S22 in the solvent extraction step S2A, the extraction residue obtained in the first step S21 is brought into contact with an alkyl sulfide and subjected to solvent extraction to extract palladium, and then back-extracted to obtain a back-extract containing palladium. This is a step of producing a product liquid and an extraction residual liquid.

Alkyl sulfide is used as the extractant. The alkyl sulfide is not particularly limited, but it is preferable to use industrially available dihexyl sulfide or dioctyl sulfide, and more preferably to use dihexyl sulfide. Furthermore, when using an industrially available analogous compound, it is necessary to determine from the viewpoint of selectivity with respect to impurity elements. Note that the alkyl sulfide is preferably used after being diluted with a hydrocarbon diluent to a concentration of, for example, 10% by volume to 50% by volume. Further, the extraction time is preferably 3 hours or more.

Here, the pH of the extraction residue obtained in the first step S21 (target of the solvent extraction in the second step S22) is not particularly limited, but the pH is between 0.5 and 2.5 prior to contact with the alkyl sulfide. It is preferable to adjust it within this range. As a result, even if gold, selenium, antimony, tin, etc. remain in the extraction residue obtained in the first step S21, crud will be generated at each processing stage of solvent extraction, washing, and back extraction. It is possible to prevent the co-extraction of impurity elements that occurs when If the pH of the extraction residue to be treated is less than 0.5, the effect of preventing co-extraction will be insufficient, and impurity elements will be likely to be co-extracted together with palladium. On the other hand, if the pH of the extraction residue exceeds 2.5, bismuth may precipitate and platinum group elements may co-precipitate.

In addition, if the remaining amount of tellurium, antimony, and tin in the extraction residue to be treated (extraction residue obtained in the first step S21) exceeds several tens of mg/L, precipitates may be present in the extraction residue. However, it is desirable to separate the generated precipitate before use.

In addition, as the pH of the extraction residue to be treated is adjusted, the solubility of diethylene glycol dibutyl ether dissolved in the extraction residue decreases to 0.00% as the acid concentration decreases. It decreases from n to ng/L to 0.01 g/L. From this, crystals of diethylene glycol dibutyl ether precipitate in the extraction residue, and can be efficiently recovered by flotation separation.

Palladium in the organic phase obtained by the solvent extraction in the second step S22 can be back-extracted using, for example, aqueous ammonia, and a back-extraction product liquid containing palladium is obtained by the back-extraction process. In order to separate and remove coexisting impurity elements, back extraction is preferably carried out after washing treatment with, for example, hydrochloric acid at a concentration of 1 mol/L to 2 mol/L. Note that the organic phase regenerated by back extraction can be used for extraction again.

From the palladium-containing back-extraction product obtained in the second step S22, palladium of a purity that can be made into a product can be recovered by a known method (palladium purification step S221). For example, diamminepalladium(II) chloride crystals with a purity of 99.9% by weight or higher (metal equivalent) can be obtained by neutralizing with hydrochloric acid.

(3rd step)
In the third step S23 in the solvent extraction step S2A, the extraction residue obtained in the second step S22 is subjected to solvent extraction by contacting with bis(2-ethylhexyl) phosphoric acid to remove organic compounds containing cationic impurity elements. This is a step of producing a phase and an extraction residue. In the third step S23, cationic impurity elements such as bismuth, copper, lead, and nickel that are not separated and removed in the second step S22 are extracted and removed.

Bis(2-ethylhexyl) phosphoric acid is used as the extractant. In the solvent extraction in the third step S23, any acidic extractant can be used in principle, but ( When using an acidic extractant (with a large pKa value), it is necessary to increase the pH for extraction of each metal ion, which results in bismuth hydrolysis and precipitation. On the other hand, if a more acidic extractant (lower pKa value) is used, back extraction becomes difficult.

Bis(2-ethylhexyl)phosphoric acid is preferably used after being diluted with a hydrocarbon diluent to a concentration of, for example, 10% by volume to 50% by volume.

The pH of the extraction residue to be treated (the extraction residue obtained in the second step S22) is not particularly limited, but is preferably 2.5 to 4.5. If the pH of the extraction residue is less than 2.5, there is a possibility that the extraction of impurity elements will be incomplete. On the other hand, if the pH of the extraction residue exceeds 4.5 and bismuth is contained, a precipitate will be generated, making it easier to form a cladding.

The method for adjusting the pH is not particularly limited, but it is preferable to use bis(2-ethylhexyl)phosphoric acid partially converted into an alkali metal salt in advance. Specifically, for example, while mixing the extraction residue obtained in the second step S22 and bis(2-ethylhexyl) phosphoric acid, an alkali of bis(2-ethylhexyl) phosphoric acid is added as a pH adjuster. Preferred is a method using metal salts. This can be done by adding alkali to adjust the pH while mixing the extraction residue obtained in the second step S22 (extraction residue to be treated) and bis(2-ethylhexyl) phosphoric acid. This is because when the extraction residual liquid contains bismuth, the alkali may directly react with the bismuth to cause precipitation of oxychlorides and the like. That is, by performing extraction using an ion exchange reaction between alkali metal ions and impurity element ions in the extractant, precipitation of the bismuth compound can be prevented.

The organic phase obtained by solvent extraction in the third step S23 is not particularly limited, but it is preferably washed with an aqueous solution containing near-neutral salts such as sodium chloride prior to back extraction. Thereby, water droplets physically dispersed and suspended in the organic phase can be collected in the aqueous phase. Note that bis(2-ethylhexyl) phosphoric acid does not extract the chlorocomplex of platinum group elements forming anions, but physically disperses and suspends the aqueous phase in the organic phase.

The back extraction treatment for the organic phase (or the washed organic phase if washed) can be performed by a known method using a strong acid solution such as hydrochloric acid, nitric acid, or sulfamic acid. Here, when bismuth or lead is included as an impurity, it is preferable to use a hydrochloric acid solution that forms a complex with these elements and can be efficiently back-extracted even at low concentrations. The concentration of hydrochloric acid used for back extraction is not particularly limited, but is preferably 0.5 mol/L to 2 mol/L. If the hydrochloric acid concentration is less than 0.5 mol/L, precipitation may occur due to bismuth hydrolysis. On the other hand, when the hydrochloric acid concentration exceeds 2 mol/L, the solubility of lead chloride decreases due to the common ion effect, and lead chloride may precipitate. Note that the organic phase regenerated by back extraction can be used for extraction again.

Note that the extraction residual liquid obtained through the treatment in such solvent extraction step S2A, that is, the extraction residual liquid obtained through the third step S23, is subjected to the next step, caking step S3.

-Second embodiment of the solvent extraction step As shown in the flow diagram of FIG. 2, the solvent extraction step S2B as the second embodiment involves bringing the leaching product into contact with diethylene glycol dibutyl ether and subjecting it to solvent extraction. The method includes a first step S21 and a second step S22 in which the extraction residual liquid obtained from the first step S21 is brought into contact with an alkyl sulfide and subjected to solvent extraction. That is, in this second embodiment, unlike the treatment in the solvent extraction step S2A of the first embodiment, the extraction residue obtained from the second step S22 is brought into contact with bis(2-ethylhexyl) phosphoric acid. This is an embodiment in which the third step S23 of subjecting the sample to solvent extraction is omitted.

The first step S21 and the second step S22 in the solvent extraction step S2B are the same as each treatment step in the solvent extraction step S2A of the first embodiment, so the description thereof will be omitted here.

On the other hand, as described above, in the solvent extraction step S2B, unlike the treatment in the solvent extraction step S2A of the first embodiment, the extraction residue obtained from the second step S22 and bis(2-ethylhexyl) phosphorus are It does not include the third step S23 of contacting with an acid and subjecting it to solvent extraction, and the extraction residual liquid obtained in the second step S22, that is, the extraction residual liquid after solvent extraction for extracting palladium, is used in the next step. It is subjected to the cake forming step S3.

Here, in the third step S23 in the solvent extraction step S2A of the first embodiment, it is possible to remove most of the bismuth contained in the extraction residual liquid obtained from the second step S22. However, depending on the amount of bismuth contained in the extraction residue, a predetermined amount of potassium chloride is added in the next step, caking step S3, to form a cake (precipitate). ) remains the same. Therefore, from the viewpoint of effectively and efficiently separating and recovering platinum, it is preferable to omit the third step S23 in the solvent extraction step S2A of the first embodiment.

Therefore, in the solvent extraction step S2B, which is the second embodiment, the third step S23 described above is not included, and the extraction residual liquid obtained through the second step S22 after performing the solvent extraction for extracting palladium is used. , and subjected to the next step, cake forming step S3.

According to such a solvent extraction step S2B, platinum can be effectively separated and recovered in the next caking step S3 while simplifying the step. In addition, when bismuth is separated by performing the solvent extraction process in the third step S23 described above, bismuth accumulates in the organic phase in the form of a precipitate. When (organic solvent) is repeatedly used, the frequency of replacing the organic phase increases, or an operation to remove the precipitate becomes necessary. In this regard, since the solvent extraction step S2B of the second embodiment does not include the third step S23, the above-mentioned sediment removal operation is unnecessary, and frequent replacement of the organic phase is also unnecessary.

[Caking process]
(Creation process of cake containing platinum)
The mutual separation method according to the present invention is characterized by performing a caking step S3 in which a predetermined amount of potassium chloride is added to the extraction residue obtained by the treatment in the solvent extraction step S2 to form a cake (precipitate). There is.

In the mutual separation method according to the present invention, most of the platinum (Pt) contained in the extraction residue can be precipitated by producing a precipitate in the caking step S3. Thereby, the recovery rate of platinum can be effectively improved.

The potassium chloride to be added may be in the form of a solid (powder) or may be in the form of a potassium chloride solution dissolved in a solvent such as water.

In the caking step S3, the amount of potassium chloride added is important, and specifically, potassium chloride is added so that the concentration in the extraction residue is 20 g/L or more. Preferably, potassium chloride is added so that the concentration in the extraction residual liquid is 40 g/L or more, and more preferably, potassium chloride is added so that the concentration in the extraction residual liquid is 60 g/L or more. .

Here, the extraction residual liquid to be treated contains platinum at a rate of about 2.0 g/L to 2.5 g/L. Generally, when potassium chloride is added to such extraction residue to form a precipitate, there is a concern that rhodium (Rh), ruthenium (Ru), etc. will also precipitate at the same time and become impurities. Surprisingly, the inventors found that such concerns were not expressed, and that Pt was preferentially precipitated and could be effectively recovered while reducing impurities.

The reason for this is thought to be as follows. That is, the precipitation reaction of platinum group elements with potassium chloride needs to take the form of K ₂ [MCl ₆ ] (M=platinum group element), and platinum group elements need to be in a tetravalent state, but platinum The valence of group elements is affected by the redox potential. In the extraction residue obtained through the above-mentioned solvent extraction step S2, since Pt is in a tetravalent state, it becomes K ₂ [PtCl ₆ ] and precipitates as a cake. On the other hand, since the trivalent ratio of Ir, Ru, and Rh in the extraction residue is high, the formation of precipitates is slow. This results in preferential precipitation of Pt in the cake.

The present inventors have also found that the above-mentioned effect is expressed when the amount of potassium chloride added to the extraction residue is 20 g/L or more in terms of concentration in the extraction residue. The addition amount of potassium chloride of 20 g/L is a large excess amount equivalent to about 20 to 25 times the reaction equivalent of Pt, so Pt, which is in a tetravalent state, is in a trivalent state. It is presumed that this element has a stronger tendency to preferentially precipitate than other platinum group elements.

As described above, in the mutual separation method according to the present invention, most of the platinum (Pt) contained in the extraction residue can be turned into a precipitate by generating a precipitate in the caking step S3. Specifically, when the Pt in the extraction residue is 100, the Pt introduced into the Pt purification treatment was less than 90 in the conventional method, but it can be increased to 95 or more.

On the other hand, the amount of potassium chloride added is preferably such that the concentration in the extraction residue is 80 g/L or less. When potassium chloride is added so that the concentration of potassium chloride in the extraction residue exceeds 80 g/L, the distribution ratio of Pt in the formed precipitate can be increased to over 95%, and the Pt remaining in the extraction residue (Pt residual Although the distribution ratio of impurity elements such as Ir, Ru, Rh, etc. into the precipitate also begins to increase.

For this reason, it is preferable to add potassium chloride in an amount such that the concentration in the extraction residue is 80 g/L or less, thereby reducing the content of impurities in the precipitate and Pt. can be selectively recovered.

(Platinum purification process in cake (precipitate))
In the mutual separation method according to the present invention, a step (platinum purification step S31) of refining Pt from the Pt-containing precipitate obtained through the above-described caking step S3 can be performed.

Specifically, the platinum-containing precipitate obtained by the treatment in the caking step S3 is separated by filtration or the like, and then purified by a known method. For example, by dissolving the precipitate using hydrochloric acid or the like and then adding ammonium chloride, ammonium hexachloroplatinate (IV) crystals having a purity of 99.9% by weight or more (metal equivalent) are produced. Note that the purification method is not limited to this.

The platinum that is purified and recovered is obtained through the caking step S3, and corresponds to 90% or more of the Pt contained in the platinum group element-containing material. In this way, the recovery rate of Pt can be improved by passing through the caking step S3.

[Oxidation neutralization process]
In the mutual separation method according to the present invention, an oxidation-neutralization step S4 is performed in which the filtrate produced by filtering and separating the precipitate after the caking step S3 is subjected to an oxidation-neutralization treatment. Specifically, in the oxidation neutralization step S4, the pH of the filtrate is adjusted to 5 to 12, an oxidizing agent is added, and the filtrate is subjected to hydrolysis to produce a precipitate containing iridium, ruthenium, and rhodium. In this oxidation neutralization step S4, by maintaining the pH of the filtrate near neutrality, ruthenium (Ru), rhodium (Rh), and iridium (Ir), which are easily hydrolyzed, are separated as hydroxide precipitates. Note that even if Pt remains in the filtrate, it can be selectively left in the aqueous solution as a soluble alkali platinate.

In the oxidation neutralization step S4, for example, a pH adjuster is added to adjust the pH of the filtrate to 5 to 12. If the pH of the filtrate is less than 5, the hydrolysis of Ru, Rh and Ir may be incomplete, while if the pH of the filtrate exceeds 12, the hydroxide precipitates of these elements will be redissolved. there is a possibility. The pH adjuster used here is not particularly limited, and for example, a water-soluble alkali can be used, and among them, it is preferable to use sodium hydroxide.

Temperature conditions are not particularly limited, but a temperature of 60° C. to 100° C. is particularly preferable because the hydrolysis reaction is accelerated as the temperature rises. If the temperature is less than 60°C, the hydrolysis reaction may be insufficient, and if it exceeds 100°C, a pressurized reaction vessel will be required, reducing the efficiency of the treatment.

The oxidation-reduction potential (ORP, based on silver/silver chloride electrode) is not particularly limited, but is preferably 100 mV to 700 mV, more preferably 200 mV to 400 mV. In the above-mentioned leaching step S1, the platinum group elements are leached in a strongly oxidizing atmosphere, so Ru, Rh, and Ir become tetravalent chloro complexes, but in order to properly precipitate these elements, the solubility is low. It is important to reliably generate tetravalent hydroxide. At this time, if the ORP is less than 100 mV, the oxidation of the platinum group elements may be insufficient and the generation of hydroxides of Ru, Rh, and Ir may be insufficient. On the other hand, when ORP exceeds 700 mV, some of the platinum group elements are oxidized to a hexavalent state and hydroxide is dissolved, and Ru is oxidized to an octavalent state, forming volatile and explosive RuO4. There is a fear.

The oxidizing agent is not particularly limited, and includes chlorine, hypochlorite, chlorite, bromine, bromate, hypobromite, peroxosulfate, etc., which act effectively in the neutral to alkaline range. can be used. Among these, it is particularly preferable to use sodium chlorite, which is easy to store, has a low self-decomposition rate during the reaction, and is inexpensive.

[Ruthenium leaching process]
(Ruthenium leaching process)
In the mutual separation method according to the present invention, a ruthenium leaching step S5 is performed in which the precipitate obtained through the oxidation neutralization step S4 is subjected to a leaching treatment to obtain a leached product solution containing ruthenium. Specifically, in the ruthenium leaching step S5, the precipitate obtained through the oxidation neutralization step S4 is leached by adding an oxidizing agent in a strong alkaline aqueous solution with a pH of 12 or higher, and a residue containing Ir and Rh and a Ru Obtain a leaching solution containing. In this oxidation neutralization step S4, Ru is leached out as sodium ruthenate (VI) by oxidation in a strong alkaline aqueous solution.

The pH of the strong alkaline aqueous solution is 12 or more, preferably 13 or more, and when sodium hydroxide is used, the NaOH concentration is more preferably 10% by weight or more. The higher the pH, the more stable sodium ruthenate (VI) becomes in the liquid. When the pH is less than 12, hardly any sodium ruthenate (VI) is produced. Note that the pH adjuster is not particularly limited, and water-soluble alkalis are used, but sodium hydroxide is preferred among them.

The oxidation-reduction potential (ORP, based on silver/silver chloride electrode) is not particularly limited, but is preferably 100 mV to 300 mV. If the ORP is less than 100 mV, oxidation of ruthenium (IV) hydroxide to sodium ruthenate (VI) will be insufficient, and leaching of ruthenium will be insufficient. On the other hand, when ORP exceeds 300 mV, self-decomposition of the oxidizing agent becomes large, which is inefficient.

The oxidizing agent is not particularly limited, and chlorine, hypochlorite, chlorite, bromine, bromate, hypobromite, peroxosulfate, etc., which act effectively in an alkaline region, may be used. I can do it. Among these, it is particularly preferable to use sodium chlorite, which is easy to store, has a low self-decomposition rate during the reaction, and is inexpensive.

The slurry concentration of the suspension is not particularly limited, but is preferably 100 g/L or less, more preferably 10 g/L to 100 g/L. The lower the slurry concentration, the higher the leaching rate, and by keeping the slurry concentration at 100 g/L or less, a leaching rate of 90% or more can usually be obtained.

In order to obtain high-purity ruthenium, it is preferable to control the leaching of platinum group elements and other impurities by intentionally keeping the ruthenium leaching rate low by adjusting conditions such as ORP, pH, and slurry concentration.

(Ruthenium purification process)
If necessary, a ruthenium purification step S51 can be performed to purify Ru from the obtained ruthenium leaching product liquid. The ruthenium purification step S51 includes, for example, a reduction step in which a reducing agent is added to the ruthenium leaching solution to obtain a ruthenium-containing precipitate, and a ruthenium crystallization step in which the precipitate is dissolved to obtain ruthenium crystals. In the ruthenium crystallization step, for example, potassium chloride or ammonium chloride is added to an aqueous solution obtained by dissolving a ruthenium-containing precipitate in hydrochloric acid to obtain Ru crystals. This makes it possible to obtain Ru crystals of such purity that they can be commercialized.

In the reduction step, sodium ruthenate (VI) in the ruthenium leaching solution is reduced by a reducing agent to produce a precipitate of ruthenium (IV) hydroxide. Note that ruthenium (IV) hydroxide precipitates when the redox potential (based on a silver/silver chloride electrode) is around 0 mV. The reducing agent is not particularly limited, but it is preferable to use a mild reducing agent such as alcohols, ketones, or sugars that can selectively reduce only ruthenium.

In the crystallization step, ruthenium(IV) hydroxide is dissolved in hydrochloric acid as hexachlororuthenium(IV) acid or its hydrated complex ion, and potassium chloride or ammonium chloride is added to form hexachlororuthenium(IV) salt. , oxopentachlororuthenate (IV), or oxotetrachlororuthenate (IV) crystals can be obtained. Note that even if a trace amount of an impurity element is contained, all of the impurity element is distributed to the mother liquor. Through such a crystallization step, ruthenium crystals with a purity of 99.9% by weight or higher (metal equivalent) can be obtained.

In addition, in order to obtain a ruthenium compound with even higher purity, recrystallization purification by treating the crystals can be performed as necessary. For recrystallization purification, for example, the obtained crystals are reduced with a weak reducing agent such as hydrazinium chloride or sulfite ion to obtain an aqueous solution of ruthenium (III) chloride, which is then re-reduced using an oxidizing agent. An oxidizing method can be used.

[Iridium extraction process]
(Extraction of iridium and purification of rhodium from extraction residue)
In the mutual separation method according to the present invention, the aqueous solution obtained by dissolving the residue (residue containing Ir and Rh) obtained through the ruthenium leaching step S5 in a hydrochloric acid solution is subjected to a solvent extraction treatment to extract Ir. Extraction step S6 is performed. Specifically, in the iridium extraction step S6, an aqueous solution containing Ir and Rh obtained by dissolving the residue in a hydrochloric acid solution is subjected to solvent extraction by contacting with tributyl phosphate, and after extracting Ir, back extraction is performed. As a result, a back-extraction product solution containing Ir and an extraction residual solution containing Rh are produced.

When dissolving the residue containing Ir and Rh with a hydrochloric acid solution, the dissolution temperature is not particularly limited, but is preferably 60°C to 100°C. By heating and dissolving in such a range, Ir can be dissolved as hexachloroiridic (IV) acid.

Further, the concentration of the hydrochloric acid solution used for dissolution is not particularly limited, but from the viewpoint of sufficiently extracting Ir as hexachloroiridic acid (IV) acid, it is preferably 3 mol/L to 7 mol/L, and 4 mol/L to 7 mol/L. More preferred.

The oxidation-reduction potential (ORP, based on silver/silver chloride electrode) of the aqueous solution containing Ir and Rh used for solvent extraction is not particularly limited, but by adding an oxidizing agent, it is preferably 700 mV to 1200 mV, more preferably 800 mV to Adjust to 1000mV. If the ORP is less than 700 mV, the hexachloroiridate (IV) acid ion becomes unstable, and some trivalent Ir may be generated, which may not be sufficiently extracted into the organic phase. On the other hand, even if ORP exceeds 1200 mV, no further extraction effect can be obtained.

The oxidizing agent is not particularly limited, and for example, chlorine, chlorate, chlorite, hypochlorite, bromate, iodate, nitric acid, etc. can be used, but among them, platinum group It is preferable to use nitric acid, which acts as a catalyst to promote the formation of chlorocomplexes of elements.

Furthermore, when Ru coexists in an aqueous solution containing Ir and Rh that is a target of solvent extraction, it is preferable to add nitrite ions to the aqueous solution. Thereby, pentachloronitrosylruthenium (III) acid can be generated, Ru can be separated into the organic phase at the same time as Ir, and the purity of Rh in the aqueous phase can be increased.

The aqueous solution used for back extraction of the organic phase containing Ir is not particularly limited, and water or a dilute acid of 1 mol/L or less can be used, but in particular, poor phase separation and hydrolysis of impurities in the organic phase can be used. From the viewpoint of preventing this, it is preferable to use an aqueous solution containing water-soluble alkali salts such as common salt.

Here, in order to more completely back-extract the elements that coexist with Ir in the organic phase, it is effective to perform reductive back-extraction using an aqueous solution containing hydrazine or its compound, or sulfite or its salts. . However, since the reducing agent suspended or dissolved in the organic phase may lower the redox potential during extraction, in such cases, the Ir and ORP control of aqueous solutions containing Rh is required.

From the extraction residue obtained by solvent extraction, that is, the extraction residue containing Rh, Rh of a purity that can be used as a product can be recovered by a known method (rhodium purification step S61). For example, sodium hexanitrorhodate (III) is obtained by adding sodium nitrite to the extraction residue, which is dissolved in warm water to remove impurities and purified, and then ammonium chloride is added to obtain hexanitrorhodium (III). III) Rh crystals with a purity of 99% by weight or more can be obtained by the method of separating and recovering ammonium acid.

(Separation of cationic impurity elements such as bismuth)
In the above-mentioned solvent extraction step S2, like the solvent extraction step S2B in the second embodiment, a third step S23, that is, a process of solvent extraction of cationic impurity elements such as bismuth, copper, lead, and nickel, is performed. If the extraction residue obtained from the second step S22 is directly subjected to the caking step S3, the residue obtained from the ruthenium leaching step S5 is dissolved in a hydrochloric acid solution in the iridium extraction step S6. The aqueous solution obtained may be subjected to a solvent extraction process to separate cationic impurity elements.

Specifically, an aqueous solution obtained by dissolving the residue in a hydrochloric acid solution is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction to separate the organic phase containing cationic impurity elements and the extracted residue. to generate liquid. As a result, cationic impurity elements such as bismuth, copper, lead, and nickel are extracted into the organic phase and removed. Note that the specific solvent extraction method is the same as the third step S23 in the solvent extraction step S2A of the first embodiment described above, and the explanation here will be omitted.

After performing a solvent extraction process to separate and remove cationic impurity elements in this manner, the resulting extraction residual liquid is subjected to an iridium extraction step S6 to extract Ir.

Further, as in the above-mentioned example, a solvent for separating cationic impurity elements from an aqueous solution obtained by dissolving the residue obtained through the ruthenium leaching step S5 in a hydrochloric acid solution prior to the treatment in the iridium extraction step S6. The present invention is not limited to performing the extraction process, and the extraction residual liquid obtained through the treatment in the iridium extraction step S6 may be subjected to a solvent extraction process to remove cationic impurity elements from the extraction residual liquid. .

Specifically, as shown in the process diagram in Figure 3, the extraction residue containing Rh obtained in the iridium extraction step is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction to extract the cationic form. It may include an impurity removal step S60 in which an organic phase containing impurity elements and an extraction residue are generated, and cationic impurity elements such as bismuth, copper, lead, and nickel are removed from the extraction residue containing Rh. Note that the details of the solvent extraction method are the same as the third step S23 in the solvent extraction step S2A of the first embodiment described above, and the explanation here will be omitted.

The extraction residue obtained through the treatment in the impurity removal step S60 is an aqueous solution containing Rh from which cationic impurity elements have been removed, and this aqueous solution is subjected to the treatment in the rhodium purification step S61 described above. With this, the purity of the rhodium compound can be further improved.

The process diagram shown in Figure 3 is the same as the process diagram shown in Figure 2, except that an impurity removal process indicated by the symbol "S60" is added, and the processing in each process is also the same, so the details will be explained here. Further explanation will be omitted.

(Iridium purification)
If necessary, an iridium purification step S62 can be performed to treat the back-extraction product liquid obtained in the iridium extraction step S6. The iridium purification step S62 includes, for example, a reduction step in which bismuth metal is added to the back-extraction product liquid and subjected to reduction to form an alloy containing a platinum group element other than Ir and an aqueous solution containing Ir, and a reduction step in which an aqueous solution containing Ir is extracted from the aqueous solution. a crystallization step to obtain crystals. In the iridium crystallization step, for example, an oxidizing agent is added to an aqueous solution containing Ir to oxidize the solution, and then potassium chloride or ammonium chloride is added to obtain Ir crystals.

The reducing agent used in the reduction step is preferably metal bismuth. Bismuth metal does not reduce Ir ions, but tends to maintain an ORP of around +300 mV, which can reliably reduce other platinum group ions. Thereby, an alloy containing a platinum group element other than Ir and an aqueous solution containing Ir can be obtained.

In the crystallization step, first, an oxidizing agent is added again to the aqueous solution containing Ir, and the redox potential (ORP, based on silver/silver chloride electrode) is adjusted to preferably 700 mV to 1000 mV, more preferably 800 mV to 1000 mV. do. As a result, hexachloroiridate (IV) acid ions necessary for crystal formation are stably produced. If the ORP is less than 700 mV, the hexachloroiridate (IV) acid ion is unstable, and there is a possibility that trivalent Ir may be partially generated. On the other hand, when the ORP exceeds 1000 mV, a small portion of lead becomes tetravalent, and lead hexachloroate crystals having the same shape as Ir may be formed and mixed.

The oxidizing agent used in the crystallization step is not particularly limited, and for example, chlorine, chlorate, chlorite, hypochlorite, bromate, iodate, nitric acid, etc. can be used.

In the crystallization step, potassium chloride or ammonium chloride is then added to the ORP-adjusted aqueous solution. This makes it possible to selectively crystallize only Ir. Thereby, crystals of hexachloroiridate with a purity of 99.9% by weight or higher (metal equivalent) can be obtained.

In addition, in order to obtain an iridium compound with even higher purity, recrystallization purification by treating the crystals can be performed as necessary. For recrystallization purification, for example, a method is used in which the obtained crystals are reduced with a weak reducing agent such as hydrazinium chloride or sulfite ion to obtain an aqueous solution of iridium (III) chloride, and this is oxidized again. be able to.

EXAMPLES Hereinafter, the present invention will be described in more detail by showing examples, but the present invention is not limited to the following examples. In the examples, metals were analyzed by ICP emission spectrometry.

[Example 1]
Using a platinum group element concentrate as a raw material, a leaching process and a solvent extraction process for the leaching product liquid were performed as shown in the flow diagram of FIG. Thereby, a solution (undiluted solution) having a Pt concentration of 2.38 g/L was obtained as an extraction residual liquid obtained through the solvent extraction step (third step).

Potassium chloride (powder) was added to 100 ml of the obtained extraction residue so that the concentration in the extraction residue was 20 g/L. Thereafter, the mixture was stirred with a stirrer at a rotation speed of 300 rpm at a liquid temperature of 25° C. to form a cake (precipitate). Filtration treatment was performed using filter paper (131) to separate the precipitate and the filtrate.

The Pt concentration in the filtrate obtained by separation was measured and found to be 0.12 g/L, which corresponds to 4.2% of the original solution. Note that the precipitate contained Pt equivalent to 90.8% of the stock solution (distribution rate of Pt to the cake: 90.8%).

[Example 2]
A test was conducted in the same manner as in Example 1, except that an aqueous potassium chloride solution was used and added so that the concentration in the extraction residue was 40 g/L.

As a result, the Pt concentration in the separated filtrate was measured and found to be 0.13 g/L, which corresponds to 4.9% of the original solution. Note that the precipitate contained Pt equivalent to 94.5% of the stock solution (distribution rate of Pt to the cake: 94.5%).

[Example 3]
A test was conducted in the same manner as in Example 1, except that an aqueous potassium chloride solution was used and added so that the concentration in the extraction residue was 60 g/L.

The Pt concentration in the filtrate obtained by separation was measured and found to be 0.11 g/L, which corresponds to 4.6% of the original solution. Note that the precipitate contained Pt equivalent to 95.4% of the stock solution (distribution rate of Pt to the cake: 95.4%).

[Example 4]
A test was conducted in the same manner as in Example 1, except that an aqueous potassium chloride solution was used and added so that the concentration in the extraction residue was 80 g/L.

As a result, the Pt concentration in the separated filtrate was measured and found to be 0.10 g/L, which corresponds to 4.2% of the stock solution. Note that the precipitate contained Pt equivalent to 95.8% of the stock solution (distribution rate of Pt to the cake: 95.8%).

[Example 5]
A test was conducted in the same manner as in Example 1, except that an aqueous potassium chloride solution was used and added so that the concentration in the extraction residue was 100 g/L.

As a result, the Pt concentration in the separated filtrate was measured and found to be 0.09 g/L, which corresponds to 3.8% of the stock solution. Note that the precipitate contained Pt equivalent to 96.2% of the stock solution (distribution rate of Pt to the cake: 96.2%).

[Example 6]
Using a platinum group element concentrate as a raw material, a leaching process and a solvent extraction process for the leaching product liquid were performed as shown in the flow diagram of FIG. Thereby, a solution (undiluted solution) having a Pt concentration of 2.04 g/L was obtained as an extraction residual liquid obtained through the solvent extraction step (second step).

Potassium chloride (powder) was added to 100 ml of the obtained extraction residue so that the concentration in the extraction residue was 20 g/L. Thereafter, the mixture was stirred with a stirrer at a rotation speed of 300 rpm at a liquid temperature of 25° C. to form a cake (precipitate). Filtration treatment was performed using filter paper (131) to separate the precipitate and the filtrate.

The Pt concentration in the filtrate obtained by separation was measured and found to be 0.19 g/L, which corresponds to 9.3% of the original solution. Note that the precipitate contained Pt equivalent to 90.7% of the stock solution (distribution rate of Pt to the cake: 90.7%).

[Example 7]
A test was conducted in the same manner as in Example 6, except that potassium chloride (powder) was added at a concentration of 40 g/L in the extraction residue.

As a result, the Pt concentration in the separated filtrate was measured and found to be 0.11 g/L, which corresponds to 5.4% of the original solution. Note that the precipitate contained Pt equivalent to 94.6% of the stock solution (distribution rate of Pt to the cake: 94.6%).

[Example 8]
The test was conducted in the same manner as in Example 6, except that potassium chloride (powder) was added at a concentration of 60 g/L in the extraction residue.

As a result, the Pt concentration in the filtrate obtained by separation was measured and found to be 0.09 g/L, which corresponds to 4.4% of the stock solution. Note that the precipitate contained Pt equivalent to 95.6% of the stock solution (distribution rate of Pt to the cake: 95.6%).

[Example 9]
The test was conducted in the same manner as in Example 6, except that potassium chloride (powder) was added at a concentration of 80 g/L in the extraction residue.

As a result, the Pt concentration in the filtrate obtained by separation was measured and found to be 0.07 g/L, which was equivalent to 3.4% of the original solution. Note that the precipitate contained Pt equivalent to 96.6% of the stock solution (distribution rate of Pt to the cake: 96.6%).

[Example 10]
A test was conducted in the same manner as in Example 6, except that potassium chloride (powder) was added at a concentration of 160 g/L in the extraction residue.

As a result, the Pt concentration in the separated filtrate was measured and found to be 0.06 g/L, which corresponds to 2.9% of the stock solution. Note that the precipitate contained Pt equivalent to 97.1% of the stock solution (distribution rate of Pt to the cake: 97.1%).

[Comparative example 1]
A test was conducted in the same manner as in Example 6, except that potassium chloride (powder) was added at a concentration of 10 g/L in the extraction residue.

As a result, the Pt concentration in the filtrate obtained by separation was measured and found to be 0.51 g/L, which was a concentration corresponding to 25.0% of the original solution. Note that the precipitate contained Pt equivalent to 75.0% of the stock solution (distribution rate of Pt to the cake: 75.0%).

Tables 1 and 2 below summarize the results of Examples and Comparative Examples.

From the results shown in Tables 1 and 2, it is clear that among the platinum group elements contained in the extraction residue, Pt is preferentially formed into a cake, and the distribution ratio to the cake is high, exceeding 90%. I understand. Furthermore, it is found that it is particularly preferable to add potassium chloride so that the concentration of potassium chloride in the extraction residue is 40 g/L or more, since the distribution ratio to the cake becomes about 95% or more and stabilizes.

On the other hand, even when potassium chloride is added to the extraction residue, if the concentration is 10 g/L (Comparative Example 1), the distribution rate to the cake is as low as 75%, and the effect of improving the Pt recovery rate is sufficient. It can be seen that this does not appear in

Furthermore, when the concentration of potassium chloride in the extraction residue exceeds 80 g/L, the effect of improving the distribution ratio of Pt to the cake gradually decreases, while elements that become impurities (Ir, The distribution rate of Ru, Rh) to the cake is increasing. From this, it can be seen that the amount of potassium chloride to be added is preferably such that the concentration of potassium chloride in the extraction residue is 80 g/L or less.

[Example 11]
The process of Example 6 was carried out according to the flowchart in FIG. 2, and the cake filtrate was subjected to an oxidation neutralization process, a ruthenium leaching process, and an iridium extraction process. Thereafter, as shown in the flow diagram of Figure 3, the extraction residue obtained from the solvent extraction in the iridium extraction step is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction to remove cationic impurities. An impurity removal process was performed to generate an organic phase containing elements and an extraction residue (impurity removal step), and an extraction residue from which cationic impurity elements were removed was obtained. In addition, the extraction residual liquid subjected to the impurity removal process (extraction residual liquid after treatment in the iridium extraction process) is referred to as "extraction residual liquid a", and the extraction residual liquid produced through the impurity removal process is referred to as "extraction residual liquid b". shall be.

Table 3 below shows the analysis results of extraction residual liquid a and extraction residual liquid b. As can be seen from the results shown in Table 3, cationic impurity elements could be significantly removed from extraction residue b obtained through the impurity removal process compared to extraction residue a before treatment. . It is considered that by subjecting the extraction residue Y obtained in this manner to a rhodium purification step (see the flow diagram of FIG. 3), the rhodium purification efficiency can be effectively improved.

Claims (5)

A method for mutually separating platinum group elements from a platinum group element-containing material containing impurity elements, the method comprising:
a leaching step of suspending the platinum group element-containing material in a hydrochloric acid solution, adding an oxidizing agent, and subjecting it to leaching to obtain a leaching product solution containing the platinum group element;
a solvent extraction step of subjecting the leaching product to a solvent extraction process;
A caking step in which potassium chloride is added to the extraction residue obtained by the solvent extraction treatment in an amount of 20 g / L or more and 80 g / L or less in the extraction residue to generate a precipitate,
The solvent extraction step includes:
A first step of contacting the leaching product solution with diethylene glycol dibutyl ether and subjecting it to solvent extraction to form an organic phase containing impurity elements and an extraction residue;
A second step of contacting the resulting extraction residue with an alkyl sulfide and subjecting it to solvent extraction to extract palladium, followed by back extraction to form a back extraction product solution containing palladium and an extraction residue;
have
A method for mutual separation of platinum group elements. The solvent extraction step includes :
A third step in which the extraction residue obtained in the second step is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction to form an organic phase containing cationic impurity elements and an extraction residue. process and
It further has
The method for mutually separating platinum group elements according to claim 1, wherein the extraction residue obtained in the third step is subjected to the caking step. moreover,
The method for mutually separating platinum group elements according to claim 1 or 2, comprising a platinum purification step of refining platinum contained in the precipitate obtained in the caking step. moreover,
The precipitate obtained in the caking step is separated by filtration, the pH of the resulting filtrate is adjusted to 5 to 12, and an oxidizing agent is added to hydrolyze it to produce a precipitate containing iridium, ruthenium, and rhodium. an oxidation neutralization step,
A ruthenium leaching step in which the precipitate obtained in the oxidation neutralization step is leached in an alkaline aqueous solution with a pH of 12 or more by adding an oxidizing agent to produce a residue containing iridium and rhodium and a ruthenium leaching product liquid;
An aqueous solution containing iridium and rhodium obtained by dissolving the residue obtained in the ruthenium leaching step in a hydrochloric acid solution is subjected to solvent extraction by contacting with tributyl phosphate to extract iridium, and then back-extracted to remove iridium. The method for mutually separating platinum group elements according to any one of claims 1 to 3 , further comprising: an iridium extraction step of producing a back-extraction product solution containing a back-extraction product solution and an extraction residual solution containing rhodium. moreover,
The precipitate obtained in the caking step is separated by filtration, the pH of the resulting filtrate is adjusted to 5 to 12, and an oxidizing agent is added to hydrolyze it to produce a precipitate containing iridium, ruthenium, and rhodium. an oxidation neutralization step,
A ruthenium leaching step in which the precipitate obtained in the oxidation neutralization step is leached in an alkaline aqueous solution with a pH of 12 or more by adding an oxidizing agent to produce a residue containing iridium and rhodium and a ruthenium leaching product liquid;
An aqueous solution containing iridium and rhodium obtained by dissolving the residue obtained in the ruthenium leaching step in a hydrochloric acid solution is subjected to solvent extraction by contacting with tributyl phosphate to extract iridium, and then back-extracted to remove iridium. an iridium extraction step that produces a back extraction product solution containing rhodium and an extraction residual solution containing rhodium;
The extraction residue obtained in the iridium extraction step is brought into contact with bis(2-ethylhexyl) phosphoric acid and subjected to solvent extraction to form an organic phase containing cationic impurity elements and an extraction residue, and remove impurities. The method for mutually separating platinum group elements according to claim 1 , further comprising a step of removing impurities.

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