JP3087758B1 - Method for recovering valuable metals from copper electrolytic slime - Google Patents

Abstract

The present invention selectively recovers gold, a platinum group element, selenium, and tellurium from copper electrolytic slime only by a simple wet operation in high yield. SOLUTION: (a) A slurry of copper electrolytic slime is treated with chlorine to leach valuable metals, and then (b) bis (2-butoxyethyl) ether is mixed with the chlorine leachate to extract gold into an organic phase. Then, gold was recovered by reduction with oxalic acid, and (c) the raffinate was mixed with trioctylmethylammonium chloride and tributyl phosphate to extract the platinum group element into the organic phase, and reduced with hydrazine and sodium hydroxide. Then, the platinum group elements are collected and separated, and (d) this raffinate is reduced with sulfur dioxide to collect selenium and tellurium in sequence.

Description

DETAILED DESCRIPTION OF THE INVENTION

[0001]

BACKGROUND OF THE INVENTION 1. Field of the Invention The present invention relates to a method for recovering valuable metals from anode slime (copper electrolytic slime) generated in an electrolytic copper refining process.

[0002]

2. Description of the Related Art Conventionally, copper electrolytic slime has been decopperized by a wet method and then selenium, antimony, lead, and the like by a dry method.
Tin, bismuth, tellurium, and the like were separated, and finally gold, silver, and platinum group alloys were obtained and processed by a wet method centering on electrolysis, thereby recovering individual noble metal elements.

However, this method requires a long period of time until the precious metal is recovered, so that the burden of interest during the residence period in the system is large, the amount of energy consumption is large, and automation is required for transporting solids in each process. However, there are problems such as difficulties in carrying out the treatment, contamination of the working environment by exhaust gas, and low adaptability to the composition of the slime and the form of the compound.

Therefore, in recent years, a processing method centering on a wet method has been gradually adopted. Numerous methods have been proposed for the wet copper electrolytic slime treatment process.These methods are roughly classified into the following two methods depending on whether a wet reduction method or a roasting method is used for selenium separation. Is done.

The first method is based on KE Sutliff et al, JO
M, August (1996), pp42-44, JE Hoffmann et al, Pr.
oceedings of COPPER 95-COBRE 95 International Conf
erence Volume III (1995), The Metallurgical Society
of CIM, pp. 41-57, JP-A-9-316559, and JP-A-9-316561.

That is, a part of tellurium and copper are leached from copper electrolytic slime using sulfuric acid and oxygen at high temperature and high pressure, and gold, platinum group elements, selenium, tellurium are extracted from the residue with hydrochloric acid and hydrogen peroxide or chlorine. Leaching. Next, bis (2-butoxyethyl) ether (hereinafter referred to as DBC) was mixed with the leachate to extract gold, and the raffinate was reduced with sulfur dioxide to remove selenium, tellurium, and platinum group elements. to recover. A mixture of selenium, tellurium, and a platinum group element is separated into selenium, tellurium, and a platinum group element by distillation in a metal state. The chlorine leaching residue leaches silver by treating it with aqueous ammonia, and silver is recovered as a powder from the leaching solution.

The second method is described in JE Hoffmann et al, HY
DROMETALLURGY '94, the Institution of Mining and
Metallurgy and the Society of Chemical Industry, C
HAPMAN & HALL (1994), pp. 69-105.

[0008] That is, the process up to the step of removing copper and removing tellurium by pressurized leaching of copper electrolytic slime with sulfuric acid is the same as the above-mentioned first method, but after that, the residue is mixed with sulfuric acid and roasted. At the same time, selenium is volatilized and the silver in the residue is converted into silver sulfate. The sulfuric acid roasting residue first leaches silver using an aqueous solution of calcium nitrate, and recovers silver metal by electrolyzing the leaching solution. The residue leached from silver leaches gold, platinum group, selenium, and remaining tellurium with hydrochloric acid and chlorine. This leachate is mixed with DBC to extract gold, but the principle is the same as in the first method. Further, the raffinate is reduced with hydrazine to recover the platinum group element and tellurium as metal powder.

As a method for recovering silver from the sulfuric acid roasting residue in the second method, a method using ammonia and a method using sodium sulfite have been proposed as in the first method. .

[0010]

According to the first method,
After adding a certain concentration of hydrochloric acid and chloride at the time of chlorine leaching, leaching with oxidizing chlorine, the composition of the slime to be leached, especially copper which produces chloride ions by reaction,
When the grade of selenium changes, the chloride concentration of the leachate changes. Therefore, when the content of these elements is greatly increased, there is a problem that the chloride concentration in the leachate increases, and the elution rate of antimony in the slime increases. The increase in the concentration of antimony in the leachate caused the generation of cladding in the extraction and cleaning steps of gold, and the insoluble antimony compound being mixed into the gold metal in the gold reduction step.

Further, since selenium is recovered immediately after recovery of gold, unless gold is completely recovered, all the gold remaining in the liquid will be mixed into selenium. Therefore, the concentration of gold in the organic phase could not be increased because it was necessary to prevent gold from remaining in the aqueous phase. In addition, since selenium is recovered in the presence of a platinum group element that is likely to co-precipitate with selenium, it is not possible to directly recover selenium with a purity that can be commercialized from the solution, and a purification step such as a distillation step is required. Was indispensable. Further, since the platinum group element is recovered only by coprecipitation of selenium and tellurium, it has been particularly difficult to recover iridium and ruthenium having a low precipitation rate with selenium and tellurium.

In the second method, since selenium dioxide volatilized during roasting is reduced immediately in a state in which smoke ash is mixed, selenium having a grade of 3N or more can hardly be recovered. In addition, since it passes through the roasting process, the platinum group elements are oxidized and inactivated, so that the leaching rate is lowered,
In particular, leaching of rhodium, iridium, and ruthenium, which are susceptible to oxidation, was difficult.

Further, a problem common to both methods is that:
Since tellurium is incompletely leached in the two steps of the copper removal step and the chlorine leaching step, there is a problem that the tellurium recovery step becomes complicated.

The present invention has been made in view of the above-mentioned problems of the conventional method, and only requires a simple wet operation from copper electrolytic slime.
An object of the present invention is to provide a method for selectively recovering gold, a platinum group element, selenium, and tellurium in a high yield.

In particular, even if the composition of the copper electrolytic slime changes, no trouble is caused after the gold extraction step, and even if the gold is not completely recovered, it can be prevented from being mixed into the selenium metal. Prevents the sediment of platinum group elements that easily settle into selenium metal, and in addition to platinum, palladium and rhodium,
It is an object of the present invention to provide a method capable of recovering iridium and ruthenium and efficiently recovering tellurium in one step.

[0016]

Means for Solving the Problems In order to achieve the above object, a method for recovering valuable metals from copper electrolytic slime provided by the present invention is characterized in that the method is performed according to the following steps (a) to (d). That is, (a) a step of leaching gold, a platinum group element, selenium, and tellurium by treating a slurry of copper electrolytic slime with chlorine, and (b) bis (2-butoxyethyl) in the obtained chlorine leachate. Extracting gold into an organic phase by mixing ether, washing the organic phase with hydrochloric acid, and then reducing it with oxalic acid to recover gold as a simple substance;
(c) A mixture of trioctylmethylammonium chloride and tributyl phosphate was mixed with the raffinate after extracting gold to extract the platinum group element into the organic phase, and after washing this organic phase with hydrochloric acid, hydrazine and (D) reducing the raffinate after extracting the platinum group element with sulfur dioxide, and sequentially collecting selenium and tellurium as a single substance by reducing with sodium hydroxide; , Is included.

In the step (a), the chloride concentration in the leaching solution at the time of chlorine leaching is maintained at 130 g / liter or less. In the step (b), bis (2) is removed from the chlorine leaching solution. When extracting gold with (-butoxyethyl) ether, the gold concentration in the organic phase is maintained at 85 g / liter or less.

In the above step (c), when washing the organic phase from which the platinum group element has been extracted, hydrochloric acid having a concentration of 0.75 mol / l or more is used, and the organic phase after reducing the platinum group element is reused. When used for extraction, hydrochloric acid is added to adjust the pH to 1 or less.

In the step (d), when the raffinate after the extraction of the platinum group element is reduced with sulfur dioxide, the oxidation-reduction potential of the liquid is maintained at 400 to 500 mV, and the liquid temperature is reduced to 50 mV.
８０80 ° C., selenium metal is suspended in advance before the reduction is started, and the raffinate is continuously supplied to the reduction vessel.
In addition, selenium is recovered by maintaining the difference between the selenium concentration at the start of reduction and the selenium concentration during continuous reduction at 5 g / liter or less. Further, tellurium is recovered by maintaining the oxidation-reduction potential of the liquid after the recovery of selenium at 290 to 380 mV.

[0020]

BEST MODE FOR CARRYING OUT THE INVENTION In the present invention, between a step (b) for extracting and recovering gold and a step (d) for reducing and recovering selenium and tellurium, a platinum group which can also extract gold is used. Element extraction process
By providing (c), coprecipitation of platinum group elements in selenium recovery in the subsequent step (d) is prevented, and it is possible to recover iridium and ruthenium, which were difficult to directly recover from copper electrolytic slime by the conventional method. Become. Further, even when the gold cannot be completely extracted in the step (b), it is possible to prevent a decrease in the overall yield.

Hereinafter, each step will be described in detail. First,
In the chlorine leaching of the copper electrolytic slime in the step (a), chloride ions are generated by the reaction between chlorine and the components in the slime, and the main factors of the chloride increase are copper and selenium. The reaction of forming chloride ions generated by copper and selenium is shown in the following chemical formulas 1 and 2.

[0022]

Embedded image Cu + Cl ₂ → CuCl ₂

[0023]

## STR2 ## Ag ₂ Se + 3Cl ₂ + 3H ₂ O → ₂ AgCl
+ H ₂ SeO ₃ + 4HCl

The copper and selenium contents in the copper electrolytic slime vary greatly depending on the mineral composition and the pretreatment method used as the raw material, and even if the initial hydrochloric acid and chloride concentrations are kept constant, the final leachate is Chloride concentration varies greatly.
Since the particularly harmful antimony increases rapidly in the subsequent process due to the increase in the chloride concentration, it was confirmed that the control of the chloride concentration in the leachate in the step (a) is very important.

That is, when the chloride concentration exceeds 130 g / l, the concentration of antimony greatly increases. Therefore, the chloride concentration in the leachate is preferably maintained at 130 g / l or less, and is preferably maintained at 110 g / litryl or less. More preferred. The chloride concentration in the leaching solution can be calculated from the selenium grade and the metallic copper grade in the raw material, and the chloride concentration in the mother liquor of the chlorine leaching slurry was monitored and this concentration exceeded the target concentration. In this case, the concentration can be adjusted by adding water to adjust the concentration.

By controlling the chloride concentration as described above, it is possible to cope with a change in the composition of the raw material, and it is not always necessary to perform the conventional copper removal before the chlorine leaching. Further, similarly to antimony, it is possible to control the elution of lead, whose solubility increases as the chloride concentration increases. When chloride is added by adding water on the way, antimony and the like are hydrolyzed by the addition of water to form a precipitate. Therefore, the addition before filtration is desirable.

The chlorine leachate was obtained from Bruce F Rimmer, Chemistr
As described in y and Industry, 19 (1974), pp. 63-66, in the next step (b), gold is extracted with bis (2-butoxyethyl) ether which is a gold extraction solvent. The co-extracted impurities are washed with hydrochloric acid and then heated with an aqueous solution of oxalic acid to back-extract gold and simultaneously reduce it to a metallic state. At the time of reduction, by adding a chloride such as sodium chloride, eutectoid of a hardly soluble chloride such as silver chloride can be prevented. Further, by using urea as a pH adjuster, reduction of gold can be promoted and co-precipitation of easily hydrolyzable impurities can be prevented.

In the gold extraction step, conventionally, the goal was to recover the concentration of gold remaining in the aqueous phase to 1 mg / liter or less.
It was controlled to 0 g / liter or less. However, in the present invention, gold can be extracted up to 0.1 mg / liter or less in the subsequent step of extracting a platinum group element, so that the amount of gold remaining in the aqueous phase can be increased, and It is possible to greatly increase the concentration of gold. in this case,
The upper limit of the gold concentration that can be extracted is a point at which the density of the organic phase and the aqueous phase from which the gold is extracted becomes equal. The maximum gold concentration in the organic phase varies depending on the specific gravity of the aqueous phase. For example, when 1.5 mol / l hydrochloric acid is used for washing the organic phase, the gold concentration can be increased to 85 g / l.
By increasing the gold concentration of the organic phase in this manner, the capacity of the organic washing and gold reduction equipment after the extraction can be reduced.

In the next step (c), the gold raffinate is
As described in Japanese Patent Application No. 10-362079, trioctylammonium chloride (hereinafter referred to as TOMAC)
And a tributyl phosphate (hereinafter referred to as TBP) mixture is used to extract a platinum group element. In this step, in particular, platinum and palladium, which are likely to co-precipitate with selenium in the subsequent step, can be extracted and removed up to 1 mg / liter or less, so that selenium contamination and loss of the platinum group element can be prevented while recovering the platinum group element early. be able to.

In the extraction of the platinum group element, a small amount of antimony is co-extracted together with the platinum group element, in addition to copper and selenium. This is because even if the chloride concentration is controlled during slime leaching, the antimony concentration in the leachate cannot be made completely zero, so that low-concentration antimony is always distributed in the aqueous phase and the organic phase. The co-extracted elements are washed and removed with hydrochloric acid. At this time, if the concentration of hydrochloric acid is too low, only antimony may hydrolyze and precipitate. Therefore, when washing the organic phase from which the platinum group element has been extracted,
It is desirable to use hydrochloric acid having a concentration of 0.75 mol / liter or more. Although the upper limit of the hydrochloric acid concentration is not particularly limited,
The higher the hydrochloric acid concentration, the lower the cleaning efficiency gradually,
3mo which is the chloride concentration of chlorine leaching solution of normal slime
More preferably, it is 1 / liter or less.

The organic phase which has undergone a washing step after extracting the platinum group element is recovered as a metal powder by a reduction treatment using hydrazine and sodium hydroxide. At this time, as shown in the following chemical formula 3, , A side reaction that changes the form of TOMAC proceeds. In Chemical Formula 3, R represents an octyl group, and NR ₄ Cl represents TOMAC.

[0032]

Embedded image NR ₄ Cl + NaOH → NR ₄ OH + NaCl

Since the NR ₄ OH produced by this side reaction is alkaline while being a solvent, if the process is repeated as it is, the antimony in the aqueous phase is hydrolyzed, and a precipitate is formed during the extraction. Therefore, hydrochloric acid is added to the organic phase after the back extraction, and the precipitate is prevented from being returned to the original form as shown in the following chemical formula 4. Completely NR ₄ O
In order for H to return to the original NR ₄ Cl, hydrochloric acid may be added so that the pH becomes 1 or less. The lower limit of the pH is not particularly limited, but is preferably pH 0 or more in consideration of economy, increase in hydrochloric acid drainage, and the like.

[0034]

Embedded image NR ₄ OH + HCl → NR ₄ Cl + H ₂ O

By previously lowering the pH of the aqueous phase,
Although a method of preventing a pH increase by OH-type TOMAC is possible, it can be said that pH adjustment of the extractant itself is a more preferable method in consideration of a local pH increase at an interface between an organic phase and an aqueous phase.

From the raffinate of the platinum group element, the following steps
In (d), selenium is recovered by reduction with sulfur dioxide. In the raffinate, platinum and palladium, which are particularly easy to precipitate among the platinum group elements, have already been separated, but rhodium, ruthenium, and iridium, which are elements that are more difficult to precipitate, remain, preventing the coprecipitation of these elements. By taking measures, the selenium quality can be further increased.

In general, an aqueous solution containing selenous acid or selenic acid is reduced at about 50 ° C. or less to produce red selenium, and platinum-group elements other than platinum and palladium hardly coprecipitate in this form of selenium. However, if this low-temperature reduction method is directly applied to a slime leachate, the reaction rate becomes extremely slow because the acid concentration is too low. On the other hand, even in a temperature region higher than this temperature, that is, in a region of 50 to 80 ° C., there is a selenium generation region in which the precipitation rate of the platinum group element is low, although not as much as red selenium. Since selenium is easily precipitated, it is considered unsuitable for industrially recovering selenium.

However, by continuously supplying the solution to the reaction tank and keeping the difference between the initial selenium concentration in the reaction tank and the selenium concentration during the continuous reduction reaction as small as possible, it is possible to reduce the amount of newly generated rubber per unit time. The ratio of selenium can be kept low. At the same time, by suspending the selenium metal powder at the initial stage of reduction, newly generated selenium precipitates on the surface of the selenium metal powder, and even if rubbery selenium precipitates, mutual fusion can be prevented. Was revealed. The larger the change in selenium concentration is, the more rubbery selenium is precipitated. However, if the fluctuation width is 5 g / l or less, improvement is observed, and if the fluctuation width is 2 g / l or less, there is no change in the selenium produced at high temperature. Reducts are collected.

Regarding the reduction potential by sulfur dioxide, 4
If it is less than 00 mV, the precipitation of impurities increases even in the temperature range of 50 to 80 ° C. Conversely, if it exceeds 500 mV, the reduction ratio of selenium decreases, so the range of 400 to 500 mV is preferable. Further, the potential for sufficiently satisfying the quality and the yield at the same time is particularly preferably in the range of 440 to 480 mV.

By the combination of these conditions, both qualities and physical properties are equivalent to selenium powder obtained by conventional sulfur dioxide reduction of selenous acid, specifically, 99.
Selenium with a quality of 9% or more can be recovered.

From the liquid obtained by reducing and recovering selenium, tellurium is recovered by further reducing the reduction potential of the liquid. When the reduction potential is less than 290 mV, the concentration of sulfur dioxide in the liquid becomes close to the saturation concentration, the efficiency of sulfur dioxide decreases, and
The reduction of coexisting copper ions proceeds, and copper (I) chloride precipitates and is mixed in tellurium, which is not preferable. Also, 38
Above 0 mV, tellurium reduction itself becomes incomplete. Therefore, the reduction potential of tellurium is 290-38
The range of 0 mV is preferable, but the range of 310 to 350 mV is more preferable in order to simultaneously sufficiently satisfy the quality and the reduction ratio of tellurium. The temperature of the liquid does not affect the tellurium grade, but the higher the temperature, the higher the reaction rate. Particularly preferred reduction temperature is 80 to 100 ° C.

Although the tellurium obtained here does not reach the quality of the product level, it can be purified by an existing method such as an electrolysis method, a precipitation method or a solvent extraction method after dissolution to obtain 9%.
Purity can be increased from 9.9 to 99.99%.

Although the present invention is directed to the treatment of copper electrolytic slime, the present invention is not limited thereto.
The same can be applied to the processing of a mixture containing either selenium or tellurium.

[0044]

【Example】Example 1  Copper electrolytic slime bag having the composition (% by weight) shown in Table 1 below
That is, 918 kg of composition A, 116 kg of composition B, and
39 kg of the resulting C were collected and mixed with 3560 liters of water.
To make a slurry.

[0045]

[Table 1]

The slurry was continuously introduced into the reaction vessel at a controlled flow rate such that the residence time in the vessel was 4 hours.
Chlorine was blown in so that a potential of V or more was maintained, and slime was leached. After chlorine leaching, the slurry was subjected to solid-liquid separation using a centrifugal separator, and the residue was washed until the washing liquid became colorless. Table 2 below shows the composition (g / g) of the leachate including the residue cleaning solution.
l) and the composition of the residue after washing (%, where Pt, Pd, I
r, Ru and Rh are ppm), and the leaching rate of each element (%)
Is shown. Incidentally, the chloride concentration in the leachate was 86.9 g / l.

[0047]

[Table 2]

2580 liters of the obtained leachate were collected and gold was extracted using 645 liters of DBC. The extraction was carried out using a mixer settler, and the organic phase and the aqueous phase were mixed in a counter-current stirrer in two continuous stages. The residence time of mixing was 10 minutes and the phase separation time was 20 minutes. The organic phase from which gold was extracted was washed with 645 liters of 1.5 mol / l hydrochloric acid using a mixer settler in three stages with continuous countercurrent. The flow of the liquid for extraction and washing is shown in FIG. 1 as a conceptual diagram of a mixer settler.

Table 3 below shows the composition (g / l) of the raffinate, the washing solution, and the extracted organic phase after washing. The extraction ratio of gold was 99.85%, and the loss of gold due to washing of the solvent from which the gold was extracted was 0.09%. The gold in the cleaning liquid can be recovered by supplying the cleaning liquid to the extraction stage.

[0050]

[Table 3]

90 liters of the organic phase from which gold was extracted were fractionated, and 50 liters of water and 2900 g of salt (1 mo)
1 / l), 1750 g of oxalic acid dihydrate (1.5 times the calculated amount)
90 times), mixed with 1110 g of urea (1.5 times the calculated amount),
The temperature was raised to ° C. The potential of this mixture was measured at 90 ° C., and when it reached 680 mV, the mixture was cooled and solid-liquid separated. The recovered gold metal was washed with methanol and then pickled with hydrochloric acid. The composition (g / l) of the organic phase after reduction and the aqueous phase after back extraction, and the grade (ppm) of reduced gold are shown in Table 4 below. The reduction rate of gold is 97.
Although the total amount of the aqueous phase and the organic phase after the reduction containing gold is repeated, the average value of the reduction ratio of gold is 99. 7%.

[0052]

[Table 4]

2580 liters of the raffinate after the gold extraction was
Extraction was performed using 2580 liters of a mixed solvent of MAC: TBP = 1: 1, and the organic phase after extraction was further reduced to 0.75 m.
Washed with 1290 liters of ol / l hydrochloric acid. The above extraction was carried out using a mixer settler in 5 extraction stages and 5 washing stages. The residence time was 10 minutes for both extraction and washing, and the phase separation time was 20 minutes. The flow of the liquid for extraction and washing is shown in FIG. 2 as a conceptual diagram of a mixer settler. Unlike the case of gold extraction, the washing liquid is finally mixed with the raffinate,
The phase ratio of the extraction stage is O / A = ２. Composition (g / l) of platinum group raffinate after mixing and platinum group extracted organic phase after washing
Are shown in Table 5 below.

[0054]

[Table 5]

In this step, the platinum group element is Pt:
99.9% or more, Pd: 99.9% or more, Ir: 99.1
%, Ru: 84.2% and Rh: 80.0% are extracted, and the gold remaining in the gold extraction step is also recovered in this step, so that the yield from the leachate reaches 99.99%.

The extracted organic phase was sampled in 50 liters, mixed with 25 liters of water, and the pH was raised to 13 with sodium hydroxide.
Reduction was performed by adding 60% hydrazine hydrate until the pressure reached mV. After maintaining this potential and pH for 1 hour, solid-liquid separation was performed. The composition of the organic phase and the aqueous phase after reduction (g / l) and the composition of the platinum group reduced product (wet%) are shown in Table 6 below. In the reduced product, Bi and Pb were detected in addition to the components shown in Table 6.

[0057]

[Table 6]

The raffinate after the extraction of the platinum group element was reduced with sulfur dioxide to recover selenium. That is, an aqueous solution containing sulfuric acid having a pH of 0.2 and selenous acid having a selenium concentration equivalent to 1 g / l was charged into a reaction vessel, and 10 g / l of metal selenium powder was suspended. And the sulfur dioxide was dissolved until the potential reached 460 mV. Then
The raffinate shown in Table 5 was continuously supplied into the reaction tank.
The feed rate was maintained such that the residence time in the vessel was 4 hours.

The composition of the mother liquor (g / l) after reduction and the grade (ppm) of the recovered selenium are shown in Table 7 below. Other than the elements described in Table 7, bismuth, antimony, lead, and arsenic, which are elements that are not reduced in principle by sulfur dioxide, are 10
ppm or less, and sulfur was 26 ppm. The reduction ratio of selenium was 91.2%. The liquid was passed continuously for 10 days, but no sticking due to the precipitation of rubber-like selenium was observed. In the particle size distribution and the results of observation with an operating electron microscope, metal selenium reduced at 85 ° C. from the liquid having the same composition was also observed. And little difference was seen.

[0060]

[Table 7]

After reducing and recovering selenium as described above, tellurium was recovered by further reducing with sulfur dioxide at a low potential. That is, the liquid after selenium reduction was heated to 85 ° C., and sulfur dioxide was blown in until the potential reached 330 mV. Thereafter, the liquid after selenium reduction was continuously supplied so that the residence time was 4 hours, and the potential and the temperature were maintained at the initial values. The composition (g / l) of the liquid after tellurium reduction and the grade (%) of tellurium metal are shown in Table 8 below. It was confirmed that of the platinum group elements remaining in the liquid, except for ruthenium, they were completely captured in tellurium and recovered.

[0062]

[Table 8]

[0063]Example 2  The leachate shown in Table 2 of Example 1 was replaced with DBC as an extractant.
Extraction of gold in two stages in countercurrent with O / A = 1/16 phase ratio
Was performed. Gold concentration in raffinate and organic phase after extraction
(G / l) is shown in Table 9 below.

[0064]

[Table 9]

When the DBC in the organic phase is increased to about 60 g / l, the gold concentration in the aqueous phase increases, but the recovery rate of gold, that is, the amount of gold in the first-stage organic phase is reduced. The value obtained by dividing by the sum of the gold amount and the gold amount of the second stage of the raffinate liquid is 98.8.
%Met. In addition, as described in Example 1, the entire amount of the remaining gold can be recovered in the next step of extracting the platinum group.

This organic phase was treated under the same conditions as in Example 1
After washing with 1.5 mol / l hydrochloric acid in three stages, reduction was carried out with oxalic acid in the presence of urea and sodium chloride. The quality (ppm) of the reduced gold is shown in Table 10 below. In this example, the load of the gold concentration in the organic phase was increased, but no remarkable increase in the impurity element was confirmed.

[0067]

[Table 10]

[0068]Comparative Example 1  Salts during leaching using copper electrolytic slime of high selenium grade
The effect of chloride concentration was investigated. That is, the copper electrode having the composition (%) in Table 11
Water is added to 300 g of the solution slime to make 1 liter.
Increase the potential of the slurry to 1000 mV or more with chlorine
Leaching. Leachate composition (g / l) and residue composition
(%) And the leaching rate (%) are shown in Table 12.

[0069]

[Table 11]

[0070]

[Table 12]

The obtained leachate was subjected to DBC treatment in the same manner as in Example 1.
, And washed. As a result, when the total amount of antimony in the leachate was 100%, 42.7% of antimony and 12.2% of arsenic precipitated in the washing step, and it was difficult to continue the extraction operation due to the formation of the precipitate. Further, even after the washing, 30.8% of antimony remained in the organic phase, and in the subsequent gold recovery step, antimony was hydrolyzed and mixed into the reduced gold.

Therefore, the copper electrolytic slime of composition A shown in Table 1 used in Example 1 and the copper electrolytic slime shown in Table 11 were mixed at a ratio of 1: 1 and chlorine leaching was carried out in the same manner as described above. . As a result, the chloride concentration was reduced to 142 g / l, but the antimony concentration was as high as 8.29 g / l.

[0073]Comparative Example 2  The ratio between the leachate and DBC described in Table 2 of Example 1 was changed.
To prepare a solvent solution having a gold concentration of 85 and 90 g / l.
Was. This solvent solution was combined with 1.5 mol / l hydrochloric acid and the phase ratio O /
When mixed at A = 1/1, the gold concentration was 90 g / l.
In, the organic phase became heavier than the aqueous phase, and the phases were inverted up and down. Ma
When the gold concentration is 85 g / l, oil-water separation is performed,
The phase separation rate became much slower. Therefore, gold in DBC
If the concentration exceeds 85 g / l, industrial implementation becomes difficult.
You.

[0074]Comparative Example 3  The raffinate shown in Table 3 produced in Example 1 and TOMAC:
A mixture of TBP = 1: 1 with a phase ratio O / A = 1/1
Extracted. The organic phase was divided into four parts, each of which was 0.2
5, 0.50, 0.75, 1.0 mol / l hydrochloric acid
And washed 5 times at a phase ratio of O / A = 1/1. Dissolution after washing
The suspension is recovered by filtering the medium, washed and dried, and then weighed.
Weighed. The results are shown in Table 13 below.

[0075]

[Table 13]

As can be seen from the results, when the hydrochloric acid concentration was less than 0.75 mol / l, the precipitation increased remarkably, and the operation with the mixer settler was difficult. Hydrochloric acid concentration is 0.
At 75 mol / l or more, no precipitation was observed in appearance, and no abnormality was observed in operation with a mixer settler.

[0077]Comparative Example 4  In Example 1, after extracting the platinum group element, reduction reverse extraction was performed.
The discharged organic phase (pH 3) is supplied to the extraction step as it is.
Caused a large amount of sediment in the mixer settler
did. Hydrochloric acid is added to this solvent to form a salt without precipitation.
The acid concentration was investigated.
It was confirmed that generation could be prevented.

[0078]Comparative Example 5  Using the raffinate after the extraction of the platinum group elements described in Table 5 of Example 1
Change the reduction potential to reduce selenium and impurities.
The movement was investigated. The reduction temperature is 65 ° C and batch reduction is performed.
gave. The result is calculated as the precipitation rate of copper, selenium, and tellurium (%).
As shown in Table 14.

[0079]

[Table 14]

From these results, it is found that selenium has a potential of 500 mV.
Exceeds 400 mV, the reduction rate is remarkably reduced, and when it is less than 400 mV, the incorporation of tellurium and copper increases rapidly.

[0081]Comparative Example 6  Using the raffinate after the extraction of the platinum group described in Table 5 of Example 1,
By changing the temperature at the time of selenium reduction in batch,
Reduction rate (%), recovered ruthenium and rhodium impurities
Table 15 shows the results obtained by examining the order (ppm).

[0082]

[Table 15]

As a result, it has been found that when the temperature is lower than 50 ° C., the impurity quality is low, but the reduction rate is extremely slow, and when the temperature is higher than 80 ° C., the coprecipitation of the platinum group element becomes extremely large, which is not practical. On the other hand, at 40 ° C., the reduction rate of selenium hardly increased even after 8 hours. The results in Table 15 are
Since this is the sampling result immediately after the potential has dropped, 6
Although the actual yield at 5 ° C. is low, it has been confirmed that a reduction rate of 90% or more can be obtained by extending the reduction time to 4 hours.

[0084]Comparative Example 7  The raffinate after the extraction of the platinum group described in Table 5 of Example 1 was used in the reaction tank.
Into the furnace, and blow in sulfur dioxide at 65 ° C to form a batch.
Was adjusted to 460 mV.
Was precipitated.

Therefore, in the case where the residence time was maintained at 4 hours and the reduction was performed while continuously supplying the stock solution, the selenium concentration (g / l) in the initial reaction tank and the selenium seed crystal were reduced. The effect of the presence or absence of (selenium metal suspension) was investigated. The results are shown in Table 16 below. The selenium concentration after the reaction was almost 1 g / l in each case. From these results, it was confirmed that selenium becomes rubbery when either the selenium seed crystal or the initial selenium concentration is not appropriate.

[0086]

[Table 16]

[0087]Comparative Example 8  In the tellurium reduction of Example 1, the residence time and temperature change
Without changing the reduction potential, only the reduction of tellurium and copper
The original rate (%) was investigated. Table 17 shows the results. Table 17
According to the result, when the potential was 290 mV or less, the precipitation rate of copper was high.
Too high, and above 370mV, the reduction rate of tellurium
It turns out that it falls remarkably.

[0088]

[Table 17]

[0089]

According to the present invention, gold, a platinum group element, selenium, and tellurium can be selectively recovered in high yield by only a simple wet operation of chlorine leaching of copper electrolytic slime. It is extremely useful industrially.

In addition, even if the composition of the copper electrolytic slime changes, no trouble is caused after the gold extraction step, and even if the gold is not completely recovered, it can be prevented from being mixed into the selenium metal. It prevents the sedimentation of platinum group elements that easily precipitate into selenium metal, and can recover iridium and ruthenium in addition to platinum, palladium and rhodium.

[Brief description of the drawings]

FIG. 1 is a conceptual diagram of a mixer settler for explaining a gold extraction step in Example 1. The solid arrows in the figure indicate the flow of the aqueous phase, the dotted arrows indicate the flow of the organic phase, and ○ indicates the mixture of the aqueous phase and the organic phase.

FIG. 2 is a conceptual diagram of a mixer settler for explaining a platinum group element extraction step in Example 1. The solid arrows in the figure indicate the flow of the aqueous phase, the dotted arrows indicate the flow of the organic phase, and ○ indicates the mixture of the aqueous phase and the organic phase.

──────────────────────────────────────────────────続 き Continuing on the front page (72) Inventor Takashi Hashikawa 3-5-3 Nishiharacho, Niihama-city, Ehime Prefecture Sumitomo Metal Mining Co., Ltd. Besshi Office (72) Inventor Takahiro Yamada 3-5 Nishiharacho, Niihama-shi, Ehime Prefecture -3 Sumitomo Metal Mining Co., Ltd. Besshi Office (72) Inventor Yoshiaki Manabe 3-5-3 Nishiharacho, Niihama-shi, Ehime Prefecture Sumitomo Metal Mining Co., Ltd. Besshi Office (56) References JP-A-60-208431 ( JP, A) JP-A-5-311264 (JP, A) JP-A-56-3630 (JP, A) JP-A-54-145321 (JP, A) (58) Fields investigated (Int. Cl. ⁷ , (DB name) C22B 11/06 C22B 61/00

Claims (7)

(57) [Claims] 1. A method for recovering valuable metals from copper electrolytic slime, comprising treating copper electrolytic slime according to the following steps (a) to (d). (a) a step of leaching gold, a platinum group element, selenium, tellurium by treating a slurry of copper electrolytic slime with chlorine, and (b) mixing bis (2-butoxyethyl) ether with the obtained chlorine leachate. Extracting gold into an organic phase, washing the organic phase with hydrochloric acid, and reducing it with oxalic acid to recover gold as a simple substance. (C) Trioctyl chloride is added to the raffinate after the extraction of gold. A mixture consisting of methylammonium and tributyl phosphate is mixed to extract the platinum group element into the organic phase, and the organic phase is washed with hydrochloric acid, then reduced with hydrazine and sodium hydroxide to collectively separate the platinum group element as a single substance. (D) a step of reducing the raffinate after extracting the platinum group element with sulfur dioxide, and sequentially recovering selenium and tellurium as a single substance. 2. The valuable metal from copper electrolytic slime according to claim 1, wherein in the step (a), the chloride concentration in the leachate at the time of chlorine leaching is maintained at 130 g / liter or less. Collection method. 3. In the step (b), the gold concentration in the organic phase is maintained at 85 g / liter or less when extracting gold from the chlorine leachate with bis (2-butoxyethyl) ether. The method for recovering valuable metals from copper electrolytic slime according to claim 1. 4. The method according to claim 1, wherein in the step (c), when washing the organic phase from which the platinum group element has been extracted, hydrochloric acid having a concentration of 0.75 mol / l or more is used.
2. The method for recovering valuable metals from copper electrolytic slime according to 1. 5. The method according to claim 1, wherein in the step (c), when the organic phase after the reduction of the platinum group element is used again for extraction, the pH is adjusted to 1 or less by adding hydrochloric acid. Or the method for recovering valuable metals from copper electrolytic slime according to 4. 6. In the step (d), when reducing the raffinate after extracting the platinum group element with sulfur dioxide, the oxidation-reduction potential of the liquid is maintained at 400 to 500 mV, and the liquid temperature is reduced to 50 mV.
８０80 ° C., selenium metal is suspended in advance before the reduction is started, and the raffinate is continuously supplied to the reduction vessel.
The selenium is recovered by maintaining the difference between the selenium concentration at the start of the reduction and the selenium concentration during the continuous reduction at 5 g / liter or less, and the valuable metal from the copper electrolytic slime according to claim 1, characterized in that: Collection method. 7. The method according to claim 1, wherein in the step (d), tellurium is recovered by maintaining the oxidation-reduction potential of the liquid after the recovery of selenium at 290 to 380 mV.
Or the method for recovering valuable metals from copper electrolytic slime according to 6.