JP2017201048A - Copper refining slag treatment method - Google Patents

Abstract

PROBLEM TO BE SOLVED: To provide a copper refining slag treatment method suitable for recovering valuable metal from copper refining slag with high efficiency, and a copper refining slag treatment method suitable for removing a harmful substance from the copper refining slag.SOLUTION: A method for recovering valuable metal from copper refining slag is provided that includes the following step 1 to step 3: a step 1 of oxidizing copper refining slag using oxygen-containing gas in the presence of flux; a step 2 of bringing a reactant after oxidizing obtained in the step 1 into contact with a Cu-Fe alloy bath; and a step 3 of subjecting the Cu-Fe alloy bath after contact obtained in the step 2 to a refining step to recover valuable metal, the valuable metal being at least one compound selected from an Fe compound, a Cu compound, an Ag compound, an Au compound, a Pt compound, an Mo compound and an Sn compound. A method for removing a harmful compound is also provided in which the harmful compound is at least one selected from an As compound, an Sb compound, a Zn compound and a Pb compound.SELECTED DRAWING: None

Description

  The present invention relates to a method for removing harmful elements contained in copper refining slag and a method for recovering valuable metals contained in copper refining slag.

  In recent years, copper grades in raw ores have been decreasing in copper smelting, and it has become increasingly important to reduce copper loss (slag loss) caused by melting and mixing metallic copper into slag. . Concentration of harmful elements such as As, Pb, Zn, and Sb contained in the slag is also regarded as a problem. Against this background, various inventions relating to slag purification technology and reduction of slag loss of metallic copper have been applied for patents by companies that handle copper smelting.

  In Patent Document 1 (Japanese Patent Laid-Open No. 11-140554), a flux is added to a smelting furnace of a copper smelting process to lower the viscosity of the slag, and at the same time, increase the activity of copper in the slag to increase the activity of the copper in the mat. On the contrary, a method of reducing the concentration of metallic copper in the slag to 0.2 wt% to 0.5 wt% while keeping the concentration of metallic copper at 60 wt% or more is disclosed.

In this application, it is disclosed that the activity coefficient of Cu ₂ O in molten slag is increased by addition of flux, and metal copper is transferred from the slag layer to the mat layer. Furthermore, it is disclosed that the metallic copper floating in the slag settles in the lower portion of the slag as the viscosity of the slag decreases, and this transitions from the slag layer to the mat layer.

  Patent Document 2 (Japanese Patent Application Laid-Open No. 2008-95127) discloses copper smelting by absorbing and removing harmful elements in copper smelting slag such as As, Pb, Zn, and Sb in molten metal copper that coexists with slag. Slag purification techniques are disclosed.

  In this application, a copper smelting slag is provided with a container such as a purification furnace, and after the molten slag is charged into the furnace, metallic copper is added. A step of immersing a lance in the molten metal copper layer formed at the bottom of the furnace, and agitating the copper smelting slag by jetting heating fuel such as heavy oil, natural gas, pulverized coal and oxygen-containing gas from the tip of the lance At the same time, oxides and copper oxides of harmful metals in the molten slag are reduced and metallized and absorbed by the molten metal copper layer staying at the bottom of the furnace to remove harmful elements from the slag. A method of removing is disclosed.

  When oxides such as As, Pb, Zn, and Sb are metallized, some of them volatilize and go out of the furnace with exhaust gas and are recovered as dust, but the most harmful As is more harmful than other harmful elements It is disclosed that a lot is absorbed by molten metal copper. It is described that the viscosity control of the slag is performed by defining the slag temperature to 1150 ° C to 1450 ° C.

Furthermore, the reducibility of the slag atmosphere stipulates that the oxygen partial pressure in the slag is 10 ^{−8 · 5} atm> PO ₂ > 10 ^{−11 · 0} atm in terms of 1400 ° C.

  In Patent Document 3 (Japanese Patent Laid-Open No. 9-87762), mainly for copper converter slag, a reducing agent is blown into copper smelting slag, and solid phase precipitation in molten slag, which is a viscosity increasing component of copper smelting slag. A method of reducing the viscosity of slag by reducing the magnetite content to 1% or less and recovering the copper content in the copper smelting slag is disclosed.

  This application describes the well-known fact that the copper content in copper converter slag is positively correlated with the magnetite content in copper converter slag, and the need to reduce the magnetite in slag for copper recovery In Japanese Patent Publication No. 45-36105, magnetite is reduced with a sulfurizing agent such as pyrite in a method in which a substance containing sulfur is blown into copper converter slag at a high speed and copper fine particles in the slag are absorbed and recovered in the sulfide layer. It states that

  Furthermore, at the Codelco, Caletones smelter in Chile, the pulverized coal was blown into the molten copper converter slag to reduce the magnetite in the slag and to recover the copper content in the slag. Yes.

  In response to these previous findings, it was stated that among the magnetite contents in slag, especially the importance of solid phase precipitated magnetite content in molten slag was newly recognized, and solid phase precipitated magnetite in molten slag A technology has been filed for reducing the content to 1% or less, and further determining that the content has reached 1% by measuring with a magnetic quantity meter. He states that this technology has reduced the copper content in copper smelting slag from about 4% to 0.8%. In the examples of this application, when reducing by blowing propane gas, comparing the case where the end point was predicted by measuring the amount of magnetic substance in the middle of processing and the case where the amount of magnetic substance was not measured in the middle of processing It states that the iron content in copper metal has changed greatly from 2.3% to 20.0%, and describes the effectiveness of controlling the solid phase precipitated magnetite content.

Patent Document 4 (Japanese Patent Application Laid-Open No. 2002-146448) discloses a method for reducing copper slag loss in a copper smelting process, and states that this reduces copper slag loss. As its characteristics, T.W. Fe / SiO ₂ (molar ratio) is defined to be 1.0 to 1.5, the Al ₂ O ₃ content is 4 to 8 wt%, the viscosity is 300 mPa · s or less, and T. Defining a Fe / SiO ₂ (molar ratio) to 1.1 to 1.4, it defines a content of _Al 2 _{O 3} and 5-7 wt-%.

  Furthermore, a method is described in which slag sampling is performed, slag component analysis is performed, viscosity is measured at 1250 ° C., and an appropriate amount of solvent is added by a regression equation separately created from the slag component and the measured viscosity value.

JP-A-11-140554 JP 2008-95127 A JP-A-9-87762 JP 2002-146448 A

  As described above, many techniques relating to slag loss reduction, copper recovery from slag, and slag purification in copper smelting have been studied. However, in these prior art documents, examination is insufficient from the viewpoint of effectively using copper smelting slag, and there is still much room for improvement.

  Copper smelting slag contains harmful metals such as As, Pb, Sb and Zn as impurities, while it contains many valuable metals such as Mo, Cu, Fe and Sn. Therefore, by appropriately treating copper smelting slag, it should be possible to improve the recovery rate of valuable metals and increase the added value of slag. In particular, the idea of recovering valuable metals such as iron, molybdenum and tin from copper smelting slag is not presented in any of the above documents.

Specifically, in the technique described in Patent Document 1, fluoride such as fluorite (CaF ₂ ), cryolite (3NaF / AlF ₃ ), and a mixture thereof is added to the slag in a smelting furnace with a concentration of 0.1%. It is characterized by adding 5 to 10 wt% to lower the viscosity of the slag.

  Therefore, the use conditions stipulate that the mat and slag are in a smelting furnace in a coexisting state. The reason for this is that, since fluoride increases the activity coefficient of copper oxide in the slag, the copper oxide migrates from the slag layer to the mat layer, and at the same time, the settling rate of metallic copper in the slag increases. This is because metal copper migrates from the layer to the matte layer.

  Therefore, it is not a method for treating slag discharged from the smelting furnace, nor is it a technique for recovering valuable metals in the slag.

  The technique described in Patent Document 2 is a technique for metallizing harmful metal oxides in slag by reduction treatment for the purpose of purifying slag and allowing molten metal copper to coexist and be absorbed by molten metal copper.

  The above technology means that a very large amount of molten copper (5 to 40 wt% with respect to slag weight ratio) is present at the bottom of the furnace, and a reducing agent is blown into the molten copper layer to cause slag in the vicinity of the molten copper layer. Causing a reduction reaction of the oxide of the metal, contacting the molten copper with the metal produced by the reduction reaction by gas bubbling from the furnace bottom, and absorbing the harmful elements metallized by the reduction reaction into the molten copper. It is a technology.

  Accordingly, it is described that there is no particular attention to the shape and type of the metallic copper to be added, and copper scrap or crude copper obtained during copper smelting may be used. That is, the technique described in Patent Document 2 has an important meaning in causing a reduction reaction in the vicinity of the molten copper layer at the bottom of the furnace.

  This is because, when the above reduction reaction occurs at a location away from the molten steel layer at the bottom of the furnace, the chance that the metal generated by the reduction reaction comes into contact with the molten steel layer is greatly reduced. Moreover, in the technique described in the document, recovery of components such as iron molybdenum and tin in the slag is not considered at all.

  In the technology described in Patent Document 3, a reducing agent is blown into copper smelting slag for copper converter slag, and the solid phase precipitated magnetite content in molten slag, which is a viscosity increasing component of copper smelting slag, is 1%. It presents a method for recovering copper in copper smelting slag by reducing it to the following.

  However, although it is stated that the copper content in the copper smelting slag can be reduced from about 4% to 0.8% by this method, 0.8% copper is still present in the slag. Moreover, the said literature does not suggest collection | recovery of components, such as iron, molybdenum, and tin, in slag.

The technique described in Patent Document 4 is not a method for treating slag itself, nor is it a method for recovering valuable metals (components such as iron, molybdenum, and tin) from slag. Then, limited to for Al ₂ O ₃ flux used to control the viscosity of the slag smelting furnace, not only describes the detailed conditions for determining the amount of Al ₂ O _3.

  In view of such conventional technology, an object of the present invention is to provide a method for removing harmful substances from copper smelting slag. Another object of the present invention is to provide a method for treating copper smelting slag suitable for recovering valuable metals from copper smelting slag with high efficiency.

  As a result of intensive studies to solve the above problems, the present inventors have found a method capable of removing harmful elements contained in copper smelting slag by combining specific oxidation-reduction processes in the presence of flux. It was. Moreover, the method of collect | recovering the valuable metal element contained in copper refining slag by carrying out similarly to this was also discovered. The present invention broadly encompasses the inventions of the embodiments shown below.

Item 1 A method for recovering valuable metals from copper smelting slag, comprising the following steps 1 to 3;
1 a step of oxidizing copper smelting slag with an oxygen-containing gas in the presence of flux;
2 The step of contacting the oxidized reactant obtained in Step 1 with the Cu—Fe alloy bath, and 3 The contacted Cu—Fe alloy bath obtained in Step 2 are subjected to a refining step, and valuable metals are recovered. Process.

Item 2 The method according to Item 1, further comprising a step 4 of bringing the contacted copper smelting slag obtained in the step 2 into contact with an Fe—C alloy bath.

Item 3 The copper smelting slag is selected from the group consisting of Cu compounds, Si compounds, Ag compounds, Au compounds, Pt compounds, Fe compounds, Mo compounds, Zn compounds, As compounds, Sb compounds, Pb compounds, and Sn compounds. Item 3. The method according to Item 1 or Item 2, comprising at least one compound.

Item 4 Any one of Items 1 to 3, wherein the valuable metal is at least one compound selected from the group consisting of an Fe compound, a Cu compound, an Ag compound, an Au compound, a Pt compound, a Mo compound, and a Sn compound. The method of crab.

Item 5 The method according to any one of Items 1 to 4, wherein the step 1 is carried out under a condition of 1200 ° C or higher.

Item 6 The method according to any one of Items 1 to 5, wherein the flux is a fluorine-containing compound.

Item 7 The method according to any one of Items 1 to 6, wherein the step 2 is performed under a condition of 1460 ° C. or lower.

Item 8 The method according to any one of Items 2 to 7, wherein the step 4 is performed under conditions of 1150 ° C to 1550 ° C.

Item 9 A method for removing harmful compounds from copper smelting slag, comprising the following step A and step B;
A: A step of oxidizing copper smelting slag with an oxygen-containing gas in the presence of a flux; and B, a step of contacting the oxidized reactant obtained in step A with a Cu—Fe alloy bath.

Item 10 The method according to Item 9, further comprising a step C in which the Cu—Fe alloy bath obtained in the step B is subjected to a refining step.

Item 11. The method according to Item 9 or 10, wherein the harmful compound is at least one selected from the group consisting of an As compound, an Sb compound, a Zn compound, and a Pb compound.

Item 12 The method according to any one of Items 9 to 11, wherein the step A is performed under a condition of 1200 ° C or higher.

Item 13 The method according to any one of Items 10 to 13, wherein the step B is performed under a condition of 1460 ° C or lower.

  According to the method of the present invention, harmful compounds can be efficiently removed from copper smelting slag. Moreover, according to the method of the present invention, valuable metal compounds can be efficiently recovered from copper smelting slag. These removal and recovery methods can provide an economically advantageous method.

Method of recovering valuable metal from copper smelting slag The method of recovering valuable metal from the copper smelting slag of the present invention includes the following steps 1 to 3.

1 a step of oxidizing copper smelting slag with an oxygen-containing gas in the presence of flux;
2 The step of contacting the oxidized reactant obtained in Step 1 with the Cu—Fe alloy bath, and 3 The contacted Cu—Fe alloy bath obtained in Step 2 are subjected to a refining step, and valuable metals are recovered. Process.

  Further, in the method for recovering valuable metals of the present invention, in order to recover the Fe component contained in the copper refining slag more efficiently, the copper refining slag and the Fe- The method of the aspect which further includes the process 4 (this process may be called a "hard reduction.") Which is made to contact C alloy is also included.

  The copper smelting slag is not particularly limited as long as it is slag generated in a copper smelting furnace such as a blast furnace, an electric furnace, a reflection furnace, and a flash smelting furnace. The copper smelting slag includes, for example, a Cu compound, Si compound, Ag compound, Au compound, Pt compound, Fe compound, Mo compound, Zn compound, As compound, Sb compound, Pb compound, or Sn compound. It is preferable that two or more kinds of such compounds are contained in the copper smelting slag.

  For example, the Cu compound is usually contained at a concentration of about 0.5 to 5% by weight, more typically about 0.6 to 1.2% by weight as a Cu component with respect to the copper smelting slag. .

  For example, the Fe compound is usually contained at a concentration of about 20 wt% to 60 wt%, more typically about 35 wt% to 45 wt% as the Fe component with respect to the copper smelting slag.

  For example, the Si compound is usually contained at a concentration of about 20% by weight to 60% by weight, more typically about 30% by weight to 40% by weight as the Si component with respect to the copper smelting slag.

  For example, the As compound is usually contained at a concentration of about 0.1 wt% to 5 wt%, more typically about 0.1 wt% to 0.5 wt% as the As component with respect to the copper smelting slag. .

  For example, the Pb compound is usually contained at a concentration of about 5% by weight or less, more typically about 1% by weight or less as a Pb component with respect to the copper smelting slag.

  For example, the Sb compound has a concentration of usually about 0.01 wt% to 0.2 wt%, more typically about 0.02 wt% to 0.06 wt% as the Sb component with respect to the copper smelting slag. Including.

  For example, the Zn compound is usually contained at a concentration of 1% by weight or less, more typically 5% by weight or less as a Zn component with respect to the copper smelting slag.

  For example, the Mo compound has a concentration of usually about 0.5 wt% to 3.0 wt%, more typically about 1.0 wt% to 2.0 wt% as the Mo component with respect to the copper smelting slag. Including.

  For example, the Sn compound is usually contained at a concentration of 0.1 wt% to 0.4 wt% as the Sn component with respect to the copper smelting slag.

  The above various compounds are included in the form of sulfides, oxides and the like. For example, the total amount of sulfide contained in the copper smelting slag can be generally about 0.5 wt% to 1.5 wt% as the S component with respect to the copper smelting slag.

  In addition, components such as Mg component, Al component, P component, Cl component, Ca component, Ti component, Cr component, Mn component, Ni component, Sr component, and Zr component may be contained in the copper smelting slag.

  The above description of “(metal element name) component” is a general term for compounds composed of metal element names, and is almost synonymous with “(metal element name) compound”. Such a description relationship applies to the description in this specification.

  Valuable metals recovered by the method of the present invention are Fe compounds, Cu compounds, Ag compounds, Au compounds, Pt compounds, Mo compounds, Sn compounds and the like. Only one of these valuable metals or a combination of two or more can be recovered.

  Various elemental components contained in such copper smelting slag can be analyzed by employing a technique using fluorescent X-rays.

Process 1
Step 1 in the method for recovering valuable metals from the copper smelting slag of the present invention is a step of oxidizing copper smelting slag using an oxygen-containing gas in the presence of a flux.

  Step 1 aims to oxidize sulfides such as CuS, FeS, and AsS among the above-mentioned compounds contained in copper smelting slag. In step 2 following step 1, a compound contained in copper smelting slag is reduced using a liquid phase Fe—Cu alloy (Fe—Cu bath) (this step is referred to as “soft reduction” in this specification). In other words, if the sulfide remains in the copper smelting slag, recovery of valuable metals and / or effective removal of harmful compounds cannot be achieved.

  Therefore, in Step 1, the purpose is to oxidize the copper smelting slag so that sulfides are not present as much as possible. From the viewpoint of thermodynamics, it is preferable from the viewpoint of recovery and / or removal of the target element that the sulfide contained in the copper smelting slag is converted into an oxide.

  The oxygen-containing gas used in step 1 is not particularly limited. For example, oxygen gas, air, etc. can be mentioned. Among them, it is preferable to use oxygen gas in view of high reactivity and low heat loss.

  Step 1 is preferably performed by providing a slag treatment furnace independent of the copper smelting furnace in which copper smelting slag is generated in order to enable removal of impurities and final adjustment of necessary components.

  Such a slag treatment furnace is not particularly limited, but for example, an LF furnace is preferably used. The LF furnace is a ladle heating furnace that has an electrode heating device and can heat slag in the ladle by arc discharge. Such a slag treatment furnace is preferably provided with a hole that can be opened and closed.

  The specific method for carrying out the oxidation reaction is not particularly limited. For example, it is carried out by blowing a jet gas, which is a mixture of oxygen-containing gas and fuel gas such as natural gas or propane gas, into a copper slag slag into a slag treatment furnace charged with copper smelting slag through a lance. Is preferred.

  Such a jet gas is preferably mixed so that oxygen becomes excessive. Since the temperature in the case of the complete combustion ratio of the mixed gas of oxygen-containing gas and natural gas is about 2400 ° C., when the jet gas is produced so that oxygen is excessive as described above, the temperature is high around 2000 ° C. A gas containing oxygen can be obtained.

  Further, after the slag treatment furnace is electrically heated and reaches a predetermined temperature, an oxygen-containing gas may be blown into the copper smelting slag so as not to lower the temperature. Therefore, by performing the temperature control as described above, it is possible to efficiently oxidize the sulfide contained in the copper smelting slag.

In step 1, FeO, which is an Fe compound, contained in the copper smelting slag, which is the raw material in step 1, is oxidized. When FeO is oxidized, oxides such as Fe ₃ O ₄ and Fe ₂ O ₃ are generated. Although the melting point of FeO is relatively low at 1377 ° C., the melting point of Fe ₃ O ₄ is 1597 ° C., and the melting point of Fe ₂ O ₃ is 1566 ° C., so the melting point of copper refining slag after the oxidation process is increased. And causes an increase in slag viscosity.

On the other hand, oxides such as Fe ₃ O ₄ and Fe ₂ O ₃ are very stable in oxidation-reduction reaction compared to FeO. If FeO and the oxides of the respective components contained in the copper smelting slag are present, a reduction reaction due to FeO occurs and each component may be metalized. On the other hand, even if oxides such as Fe ₃ O ₄ and Fe ₂ O ₃ and each component contained in the copper refining slag are present, the reduction reaction of each component is difficult to occur.

Therefore, it is preferable to perform the oxidation reaction in step 1 so that Fe ₃ O ₄ , Fe ₂ O _{3, and the} like are generated. However, as described above, it is necessary to cope with an increase in melting point and an increase in slag viscosity due to Fe ₃ O ₄ , Fe ₂ O _{3 and the} like.

In order to cope with an increase in the melting point of the copper smelting slag, an increase in the viscosity of the slag, etc., it is necessary to carry out the oxidation step of step 1 in the presence of flux. Such a flux is not particularly limited as long as it is a so-called flux used in fields such as metal refining. For _example, mention may be made of _{K 2 O, Na 2 O,} Al 2 O 3, MgO, CaO, NaF, KF, CaF 2, cryolite and the like.

The temperature of the oxidation reaction in step 1 is not particularly limited. Usually, it can be carried out at a temperature of about 1200 ° C. or higher. However, among the above fluxes, K ₂ O, KF, Na ₂ O, NaF, cryolite and the like containing potassium and sodium are used as copper refining slag in a temperature range of 1400 ° C. or lower in the presence of sulfide. If added, the sulfide and the flux react before the sulfide undergoes an oxidation reaction, and tend to form fine particles such as metal Cu, metal As, and metal Sb. Since such metal fine particles are difficult to recover and / or remove, they are not preferable for the method of the present invention.

  Therefore, in order to prevent the generation of the metal fine particles, the oxidation reaction is preferably performed at a temperature of about 1400 ° C. or higher, and more preferably about 1450 ° C. or higher. The upper limit of the oxidation reaction in step 1 is preferably carried out under conditions of 1600 ° C. or less, more preferably about 1550 ° C. or less, most preferably about 1500 ° C., from the viewpoint of protecting the furnace body or economically. It is as follows.

  The amount of flux added in the oxidation step is, for example, an amount in a range where the slag viscosity can be lowered so that slag handling is possible even when FeO in the slag is 1% by weight or less of the whole. There is no limitation. Further, from the state diagram of the flux to be used, a specific addition amount can be calculated based on the blending ratio at which the melting point is lowest.

In view of these, the weight ratio can be generally set to [amount of SiO ₂ that can be present in the copper refining slag after the oxidation treatment]: [amount of flux] = 95 to 60: 40-5. In view of the conversion efficiency of sulfides in copper refining slag to oxides, [amount of SiO ₂ that can be present in copper refining slag after oxidation treatment]: [amount of flux] = 70 to 60: about 40 to 30 It is preferable to set it as the weight ratio.

  The timing of adding these fluxes to the copper smelting slag is not particularly limited. For example, in view of the generation of the above metal fine particles, it can be added usually when the copper smelting slag being heated reaches about 1400 ° C.

  Process 1 is a process in which the oxidation reaction of the sulfide contained in the copper smelting slag, which is the main purpose, may be completed. However, since process 2 following process 1 is a reduction process, it prepares for an endothermic reaction by the reduction process. Thus, it is preferable to perform step 2 while maintaining a certain amount of heat.

  The amount of heat specifically maintained is not particularly limited. For example, it is preferable that the copper refining slag after the oxidation treatment has a heat amount that normally maintains a temperature of about 1200 ° C. to 1550 ° C. As a specific method for maintaining the calorific value of the copper smelting slag, the calorific value can be maintained by adjusting the electric heating, the treatment temperature by jet gas, or the time.

  According to step 1, the compounds contained in the copper smelting slag are copper oxide, silicon oxide, silver oxide, gold oxide, platinum oxide, iron oxide, zinc oxide, arsenic oxide, antimony oxide, molybdenum oxide, lead oxide, and oxidation. Converted to oxides such as tin. The total amount of sulfide in the copper smelting slag after step 1 can usually be about 0.01 wt% to 0.02 wt% as the S component with respect to the copper smelting slag.

Process 2
Step 2 in the method for recovering valuable metals from the copper smelting slag of the present invention is a step (soft reduction) in which the oxidized reactant obtained in Step 1 is brought into contact with the Cu—Fe alloy bath.

  The Fe content in a specific Cu—Fe alloy bath is not particularly limited. For example, the amount of Fe with respect to the Cu—Fe alloy bath is usually about 5% by weight or more, more preferably about 10% by weight or more, and most preferably about 20% by weight or more.

  The Fe content in the Cu—Fe alloy bath is preferably as high as possible, but it is preferable to be comprehensively determined from an economical viewpoint.

The atmosphere in which the soft reduction is performed is not particularly limited, but may be performed in an atmosphere having a predetermined oxygen partial pressure and is not particularly limited. For example, Fe + 1 / 2O ₂ = FeO can be made equal to or lower than the equilibrium oxygen partial pressure shown by the Ellingham diagram at the above processing temperature. Specifically, under the temperature condition of 1700K, it may be normally performed in an atmosphere having an oxygen partial pressure of about 1 × 10 ⁻¹⁰ atm or less, and under the temperature condition of 1400K, oxygen of usually about 1 × 10 ⁻¹³ atm or less What is necessary is just to implement in the atmosphere of partial pressure.

More preferably, it can be below the equilibrium oxygen partial pressure of C + 1 / 2O ₂ = CO. Specifically, under the temperature condition of 1700K, it may be usually performed in an atmosphere having an oxygen partial pressure of about 1 × 10 ⁻¹⁵ atm or less, and under the temperature condition of 1400K, oxygen of usually about 1 × 10 ⁻¹⁸ atm or less. What is necessary is just to implement in the atmosphere of partial pressure.

  As a means for adjusting the oxygen partial pressure, for example, a means for covering the upper part of the reaction vessel with graphite (typically a lid made of graphite) or the like to prevent the ingress of air can be mentioned. With such means, even if air enters the reaction vessel, it can be adjusted to react with graphite to become CO gas. Similarly, a means of blowing CO gas into the reaction vessel can be employed.

  The soft reduction is performed by bringing the copper refining slag obtained in Step 1 into contact with the Cu-Fe alloy bath, and then blowing nitrogen into this or stirring (typically using an alumina pipe). The reduction reaction can be promoted by agitation, agitation using a ceramic agitation blade, or the like. In addition, even if it is the method of leaving still, sufficient reduction | restoration reaction can be promoted by a thermal convection.

  According to step 2, oxides such as copper oxide, molybdenum oxide, silver oxide, gold oxide, platinum oxide, arsenic oxide, antimony oxide, lead oxide and tin oxide contained in the copper smelting slag after step 1 are liquid. It is reduced by the metallic Fe in the phase Cu-Fe alloy and moves to the Cu-Fe alloy bath.

  And since harmful compounds, such as an arsenic oxide, antimony oxide, and lead oxide, can be removed from copper smelting slag, the method containing the said process 1 and the process 2 is made into the removal method of the harmful compound of this invention mentioned later. You can also.

  In addition, the metal Fe in the Cu—Fe alloy bath is converted to FeO by the above reduction reaction, and moves into the copper refining slag.

  The temperature conditions for performing step 2 are not particularly limited. Specifically, among the compounds oxidized in step 1, it is preferable to carry out under conditions of about 1460 ° C. or lower in order to reduce molybdenum oxide and transfer it to a Cu—Fe alloy bath efficiently. The lower limit of step 2 is not particularly limited. For example, the temperature may be a temperature at which the Cu—Fe alloy is kept in a liquid state, and it may be normally about 1200 ° C.

  After step 2, the Cu—Fe alloy bath and the copper smelting slag can be easily separated. For the subsequent step 3, either the Cu—Fe alloy bath or the copper smelting slag may be moved to another reaction vessel.

  Although a means of using metal Fe instead of the Cu—Fe alloy bath in step 2 is also conceivable, in order to perform the reduction reaction in the liquid phase, it is necessary to set the temperature to 1535 ° C., which is not a practical means. If it is a Cu—Fe alloy bath as in the present invention, the liquid is maintained even at around 1450 ° C. The Fe content at this time can be about 30 parts by weight with respect to the Cu—Fe alloy bath. Therefore, it can be said that it is economically advantageous to use the Cu—Fe alloy bath in Step 2 following the oxidation step (Step 1).

  Further, when liquid phase metal Fe is used, it is dissolved in metal Fe when subjected to the refining process (typically, electrolysis method) of step 3 required for recovery of valuable metals and / or removal of harmful compounds. It is necessary to refine the ingredients. At this time, when an electrolysis method, which is a typical example of the refining process, is adopted, very large energy is required.

  On the other hand, by using a Cu—Fe alloy bath as in the present invention, various components oxidized in step 1 are reduced by Fe in the alloy and transferred into the alloy. If this is subjected to the refining process of step 3, particularly the electrolysis method, it is not necessary to refining Fe, which requires a large amount of energy. Become. Therefore, it can be said that it is economically advantageous to use the Cu—Fe alloy bath in Step 2 following the oxidation step (Step 1).

  As an alloy combined with Fe, there is a possibility of using a component such as Si in addition to Cu, but it can be said that it is most economically advantageous to use a Cu—Fe alloy bath from the above viewpoint.

Process 3
Step 3 in the method of recovering valuable metals from the copper smelting slag of the present invention is a step of recovering valuable metals by using the contacted Cu-Fe alloy bath obtained in step 2 for the refining step.

  Examples of valuable metals recovered in step 3 include a Cu compound, an Ag compound, an Au compound, a Pt compound, a Mo compound, and a Sn compound. These compounds can also be recovered as metal elements, respectively.

  The refining process of the process 3 is not specifically limited, A well-known means can be employ | adopted. Specific examples of the refining means include wet refining and dry refining. Examples of wet refining include electrolysis and precipitation separation. Examples of dry refining include a melt refining method, a molten salt refining method, a gas-solid reaction refining method, and a volatile refining method.

  Here, the electrolysis method is a method for refining a desired metal based on a difference in ionization tendency. For example, when refining a Cu compound using this means, the Cu—Fe alloy bath obtained in step 2 is solidified to form so-called crude copper, which is then dissolved in sulfuric acid, and this is a cathode made of pure copper. And an anode made of crude copper containing impurities, and adopting a method of applying a current between the anode and the cathode, metal Cu is accumulated on the cathode. Metal Cu can be recovered by such a method.

  Moreover, regarding the recovery of Cu, when the Cu—Fe alloy bath obtained in step 2 during the refining step of step 3 is set to about 1200 ° C. or less, the solubility of the Fe component contained in the Cu—Fe alloy bath tends to decrease. Since the specific gravity of the Fe component is lighter than that of the Cu—Fe alloy bath, when the Cu—Fe alloy bath is brought to the vicinity of the above temperature, the Fe component floats on the upper part, and conversely, the Cu component is concentrated on the lower part.

  When the Cu component concentrated in this manner is recovered and subjected to a refining process such as the above-described electrolysis method, metal Cu can be obtained relatively easily.

  Further, a residue (anode mud) containing valuable metal compounds other than Cu components and harmful compounds is deposited in the lower part of the electrolysis tank. By further subjecting this residue to the above-mentioned refining process, it is possible to recover desired valuable metals or remove harmful compounds from the system.

  The refining means in step 3 may be carried out by repeating the above refining means or combining a plurality of refining means in view of the type and purity of the compound desired to be recovered; the degree of removal of harmful compounds, etc. it can.

  The valuable metal recovered in step 3 is a valuable metal contained in the Cu—Fe alloy bath after contact obtained in step 2 as described above. In this, although a small amount of Fe component is contained, it is contained more significantly in the form of iron oxide (FeO) in the copper smelting slag obtained after Step 2.

  Therefore, in order to recover the Fe component, which is a valuable metal, as a raw material for iron making, the method according to the aspect including the step 4 in which the copper smelting slag obtained in the step 2 is brought into contact with the Fe—C alloy bath is used. It can also be included in the recovery method.

Step 4 Step 4 in the method for recovering a valuable metal of the present invention is a step (hard reduction) in which the copper smelting slag obtained in Step 2 is brought into contact with the Fe—C alloy bath. By this step 4, the Fe component in the iron oxide present in a large amount in the copper smelting slag is transferred to the Fe—C alloy bath, and the Fe—C alloy bath before contact as a pig iron bath that can be used mainly as a steelmaking raw material. It can be collected together.

  In Step 4, the C component contained in the Fe—C alloy is released as CO to the outside of the system. Therefore, if necessary, the C component may be positively replenished in the Fe—C alloy bath after contact. For a specific C component replenishment method, replenishment amount, replenishment amount, and the like, known methods can be referred to.

  The content of C contained in the Fe—C alloy bath before contact is not particularly limited. Specifically, it may be generally about 4.0 wt% to 5.0 wt% with respect to the Fe—C alloy bath. In view of the processing temperature, it is preferable to be about 4.5% by weight with respect to the Fe—C alloy bath.

  Since it is preferable to separate the Cu—Fe alloy bath and the copper refining slag after the above step 2 and implement the step 3 without changing the reaction tank used in the step 2, the step 4 is separated as described above. Copper refining slag can also be transferred to a new reaction tank. In Step 4, since it may be carried out while adding the C component as described above, it is preferable that a charging hole is provided for the reaction tank together with a heating device capable of controlling the temperature within the following range. .

  Step 4 can be carried out at a temperature of about 1150 ° C. or higher, which is the temperature at which the Fe—C alloy bath keeps liquid, and the upper limit can be usually carried out at about 1550 ° C.

  Further, the Zn compound contained in the copper smelting slag is oxidized to ZnO in the above step 1. This ZnO is reduced to Zn after performing Step 4 and exists in a dissolved state in the Fe—C alloy bath (pig iron bath). Even if such pig iron in which Zn is dissolved is used as it is in the refining process as a raw material for making iron, Zn is vaporized as a gas during the refining process. There is no problem because Zn is dissolved in pig iron after performing Step 4. The content of the Fe component (mainly metal Fe) in the Fe—C alloy bath (pig iron bath) obtained in step 4 is usually 95% by weight or more based on the pig iron bath. In addition, since harmful compounds such as As, Pb, and Sb contained in the copper smelting slag in Step 2 are removed from the copper smelting slag together with the Cu-Fe alloy bath, the pig iron bath obtained in Step 4 is It can be suitably used as an iron-making material as it is.

Further, the copper smelting slag obtained in the step 4 has a harmful compound removed as described above. When the flux component added in the step 1 is removed, the main component in the copper smelting slag is silicon oxide (SiO 2). ₂ ). The copper smelting slag obtained in the step 4 can be used for a roadbed material or the like.

Method for removing harmful compounds from copper smelting slag The method for removing harmful compounds from copper smelting slag of the present invention comprises the following step A and step B;
A A step of oxidizing copper smelting slag using an oxygen-containing gas, and B A step of contacting the oxidized reactant obtained in step A with a Cu-Fe alloy.

The copper refining slag can be the same as the method for recovering valuable metals from the copper refining slag .

  The harmful compound removed by the above method is not particularly limited. For example, As compound, Sb compound, Zn compound, Pb compound, etc. can be mentioned. The valuable metal to be removed may be one of the above various compounds or two or more kinds.

Process A
Step A in the method for removing harmful compounds from the copper smelting slag of the present invention is a step of oxidizing copper smelting slag using an oxygen-containing gas in the presence of a flux.

Step A may be performed in the same manner as Step 1 of the method for recovering valuable metals from the copper smelting slag .

Process B
Step B in the method for removing harmful compounds from the copper smelting slag of the present invention is a step of bringing the oxidized reactant obtained in Step 1 into contact with the Cu-Fe alloy bath.

Step B may be performed in the same manner as Step 2 of the method for recovering valuable metals from the copper smelting slag .

In addition, since the harmful compound oxidized in the above-mentioned step A is transferred from the copper refining slag to the Cu-Fe alloy bath by the step B, it is removed from the copper refining slag, but the Cu-Fe alloy bath In order to completely remove harmful compounds from the system including the above, it is possible to refer to step 3 of the method for recovering valuable metals from the copper refining slag .

  Therefore, for the purpose of completely removing harmful compounds from the system, the method of the aspect including the step C in which the Cu—Fe alloy bath obtained in the step B is subjected to the refining step is used as the method for removing the harmful compounds of the present invention. Can also be included.

  Examples for explaining the present invention in more detail are shown below. Needless to say, the present invention is not limited to the following examples.

<Example 1>
(Copper refining slag)
The analysis values of each element of the copper smelting slag used in this example are shown in Table 1 below. In addition, the following analysis used the IPC analysis method, and the unit in a table | surface is a weight part with respect to copper smelting slag.

(About various processing temperatures)
The oxidation treatment temperature was raised to 1450 ° C. by putting copper smelting slag at room temperature into a high-purity alumina crucible and heating in an electric furnace. The reason for setting the treatment temperature to 1450 ° C. is that if it is 1400 ° C. or less, there is a risk that the sulfide and Na ₂ O react to form a single element (metallization). When the temperature is increased by mixing slag and Na ₂ O flux at room temperature, the reaction between Na ₂ O and sulfide occurs in the temperature rising process to produce metallic copper, and after the metallic copper is produced, oxygen gas is blown to oxidize it However, there is a possibility that a phenomenon may occur in which the oxide does not fully oxidize, remains in the slag until the end, and moves into pig iron only after hard reduction. Therefore, a predetermined amount of flux was added after reaching the target processing temperature of 1450 ° C. Oxygen gas was blown into the slag using a high purity alumina tube.

  The following soft reduction treatment temperature is not preferably raised to 1453 ° C. or higher from the viewpoint of Mo recovery. Therefore, it was set to 1450 degreeC.

The following hard reduction treatment temperature was 1200 ° C. in order to confirm the presence of CuFe ₂ O ₄ . In other words, it is proved that CuFe ₂ O ₄ was not present if the hard reduction treatment was performed at 1200 ° C. and the Cu ₂ O concentration in the slag after the hard reduction was sufficiently lowered from the Cu ₂ O concentration after the soft reduction. Become.

(Determination of flux amount)
-Determination of the amount of Na ₂ O An attempt was made to determine the amount of flux to be added. Table 2 below shows the element content after oxidation treatment according to the ratio of SiO ₂ and Na ₂ O. The analytical value after the oxidation reaction is also shown. Elemental analysis was measured by IPC analysis. Moreover, the unit in a table | surface represents weight%.

In the oxidation treatment step, the flow of slag deteriorates with the formation of Fe ₂ O ₃ and the oxidation treatment is not performed smoothly, so the reduction of S stops. As shown in Table _{2, SiO 2: Na 2 O} = 70: 30 By using (molar ratio) or more Na ₂ O, sulfides in the slag is sufficiently oxides reduction.

However, the concentration of Fe, Cu, etc. decreases with the increase in the amount of Na ₂ O used, but this is presumed to be due to the particularity of this experiment. In this experiment, as shown below, a single electric furnace was used, and a high-purity alumina Tamman tube was used as the crucible.

  Therefore, every time one process is performed, the slag cooled in the electric furnace and cooled the next day is collected and used as an analysis sample or a raw material for the next process. This concentration change may occur due to the occurrence of uneven distribution due to the difference in specific gravity during cooling, the occurrence of precipitation during solidification that has not been confirmed, or the penetration into the alumina crucible wall. It was inferred. These elements have nowhere to go outside the crucible.

-Determination of the amount of NaF Similarly to the above, from the viewpoint of the hard reduction of the present invention (see below for the procedure), an experiment was conducted to confirm the amount of NaF used. The amount of NaF used and the analysis result are shown together in Table 3 below. Elemental analysis was measured by IPC analysis. Moreover, the unit in a table | surface represents weight%.

In the hard reduction treatment, FeO is reduced and escapes from the slag, so that SiO ₂ becomes the main component as the reaction proceeds. For this reason, the melting point of the slag rises and a three-dimensional network network formed of SiO ₂ is formed, so that the viscosity of the slag increases and the treatment becomes impossible.

In the case of hard reduction, it is not enough to lower the melting point of SiO ₂ with Na ₂ O, and a fluoride flux that destroys the three-dimensional network network with SiO ₂ is required. Therefore, NaF was selected and the required amount was investigated. As shown in Table 3, it was necessary to add NaF in a weight ratio of 8 wt% or more with respect to SiO ₂ .

・ Confirmation of the difference due to the stirring method in the soft reduction process The removal of harmful elements by the soft reduction was mainly investigated. The soft reduction temperature is 1350 ° C. As the stirring method, N ₂ gas bubbling, standing without stirring, and stirring with SUS blades were tried.

  The results are shown in Table 4 below. Elemental analysis was measured by IPC analysis. The unit in the table is% by weight.

In the flux, Na ₂ O is added at a ratio of SiO ₂ : Na ₂ O = 70: 30. Of the above three methods, it was confirmed that any method can remove the target harmful elements As and Pb to a satisfactory level. The N ₂ gas used here is a high-purity N ₂ gas that is completely deoxygenated through metal aluminum powder heated to 600 ° C. in an alumina tube.

  However, as a problem, the recovery of Cu by soft reduction was insufficient. In any stirring method, Cu in the slag after the soft reduction did not decrease to about 0.25% or less. After that, soft reduction was repeated mainly with SUS blade stirring, but Cu after the treatment stopped at around 0.25%. Therefore, based on the hypothesis that the reduction reaction by Fe in the SUS blade occurs and the fine metal Cu particles produced thereby suspend in the slag, the Cu concentration does not decrease, and the stirring is performed with a high purity alumina tube. Was changed to manual intermittent stirring using

<Example 2>
(Copper refining slag)
The analysis values of the copper smelting slag by the raw material X-ray fluorescence analyzer manufactured by Rigaku are shown in Table 1 below.

(Oxidation process)
189 g of copper smelting slag is put into a high-purity alumina Tamman tube (Nikkato T7) having an outer diameter φ50 × inner diameter φ40 × length 150 mm, and heated to a processing temperature in an electric furnace to be melted. Hereinafter, this crucible is referred to as a Tamman tube. In order to prevent molten slag from flowing out into the electric furnace when it breaks, put the slag into a magnesia crucible (outside diameter 125mm x inside diameter 110mm x 150mm in length). Filled with refractory. A magnesia crucible supporting a tamman tube was placed in an electric furnace (HLF2030, manufactured by Hiroki) and heated to 1450 ° C. at 8 ° C./min. At 1450 ° C., 52 g of Na ₂ CO ₃ (corresponding to 30 g of Na ₂ O) and 10 g of NaF were added little by little to the slag.

The amount of Na ₂ O added is targeted at an amount of Na ₂ O: SiO ₂ = 1: 2 (molar ratio). Thereafter, a high-purity alumina tube (outer diameter φ6 × inner diameter φ4 × length 1000 mm) was inserted to the bottom of the molten slag, and oxygen gas was aerated at 500 ml / min for 1 hour. Since slag forming occurred, the alumina tube was inserted and removed and controlled so as not to overflow. Thereafter, oxygen was continuously blown, and an oxidation process for a total of 60 minutes was performed.

  Thereafter, the electric furnace is allowed to cool overnight, and slag is collected at a later date. When cryolite was used, the above NaF was changed to cryolite and added at the same ratio.

However, since there was concern about the decrease in the concentration of elements with high specific gravity such as Fe ₂ O ₃ , Cu ₂ O, MoO ₃ after the oxidation treatment, the Tamman tube was extracted from the electric furnace after the oxidation process and immediately put into water. We tried to prevent the concentration gradient from occurring in the slag due to the difference in specific gravity.

-Change in slag composition by addition of flux in oxidation treatment process Increase of slag weight by addition of flux, for example, when flux is added at a ratio of (SiO ₂ : Na ₂ O: NaF = 62: 29: 9) during oxidation treatment The influence of changes in the concentration of each component due to the following is shown below.
・ Increase of slag weight by addition of Na ₂ O W1 = 0.313W (SiO ₂ amount) ÷ 62 × 29 = 0.146W
・ Increase of slag weight by addition of NaF W2 = 0.313 W (SiO ₂ amount) ÷ 62 × 9 = 0.0454 W
Oxygen weight increment W3 = W × 0.535 ÷ 55.8 (mol number of FeO) × 1/2 × 16 = 0.076W = 0.076W where FeO becomes Fe ₂ O ₃
・ Increase of slag weight due to increase of Al2O3 due to dissolution of Tamman tube W4 ≒ 0.10W
From the above, the total slag weight W4 after the oxidation treatment is W4 = W + W1 + W2 + W3 + W4.
= W (1 + 0.146 + 0.0454 + 0.076 + 0.1)
= 1.367W
And can be calculated.

  Therefore, the concentration of each component is reduced by about 37% by adding flux in the oxidation process.

  In this example, a high-purity alumina Tamman tube was used. However, since a considerable amount is dissolved in the slag due to the use of flux, the refractory in the reaction vessel is an important point in the actual machine. . In the actual machine, the area corresponding to the slag layer will use slag self-coating technology with water-cooled panels.

(Soft reduction process)
Add 300 g of Cu to the Tamman tube. The Tamman tube was put into a magnesia crucible (outer diameter φ125 × inner diameter φ110 × length 150 mm), and coke powder was packed in the gap between the Tamman tube and the magnesia crucible. It was put in an electric furnace, heated up to 1450 ° C., 100 g of electrolytic iron was added and allowed to stand for 30 minutes, and it was confirmed that metal Fe was dissolved, and a Cu—Fe alloy bath was made. Thereafter, 100 g of oxidized slag was added to the Cu—Fe alloy bath, and the graphite lid was applied to make the atmosphere in the crucible a CO atmosphere, which was manually stirred in an alumina tube every 5 minutes for 1 minute. This operation was carried out for 60 minutes and allowed to stand for another 60 minutes, and then the electric furnace was cooled overnight, and slag was collected at a later date.
Here, the reason for using a graphite lid or surrounding the Tamman tube with coke powder is to lower the oxygen partial pressure of the entire soft reduction treatment system.

(Hard reduction process)
The graphite crucible was mixed with 673 g of electrolytic iron powder and 27 g of graphite powder, and the graphite crucible was put in a magnesia crucible and placed in an electric furnace, and heated to 1450 ° C. to produce carbon-saturated molten iron. Thereafter, the temperature was lowered to 1180 ° C., a graphite impeller was inserted into the crucible and preheated for 30 minutes, and then the furnace temperature was set to 1200 ° C. and the temperature was raised.

  Hold the furnace for 30 minutes after the furnace temperature reaches 1200 ° C, insert the preheated stirring blade into the molten iron and rotate it at 60 rpm, and add 10g / time of slag after soft reduction 6 times at 10 minute intervals did. After stirring for 60 minutes, the mixture was allowed to stand for 60 minutes, and the electric furnace was allowed to cool overnight, and then slag was collected at a later date.

  Here, the temperature is raised to 1450 ° C. in order to accelerate dissolution of electrolytic iron. Thereafter, the temperature is lowered to 1180 ° C. to guarantee that the processing temperature never exceeds 1210 ° C.

(Experimental result)
Tables 6 to 8 below show the results of Example 2. The result of having analyzed the copper refining slag after the completion | finish of each process with the fluorescent X ray analyzer, respectively is shown. In addition, a numerical value is weight% with respect to copper smelting slag.

(Regarding the oxidation process)
Regardless of the type of fluoride flux, the target sulfide was oxidized (with regard to the soft reduction process)
It was confirmed that almost no sulfide was present in the slag after the oxidation treatment, and the following was revealed as a result of the soft reduction treatment of the slag after the oxidation treatment.

  It was confirmed that harmful elements such as As and Pb present in the slag after the soft reduction treatment are sufficiently removed by the soft reduction treatment. It was also found that if it is a fluoride-based flux, it does not depend on the type.

It became clear that the lower oxygen partial pressure in the atmosphere in the soft reduction treatment promotes the reduction of CuO. Further, it is clear that the use of the graphite lid has a lower Cu ₂ O concentration after the soft reduction than in the air or in a high purity nitrogen gas atmosphere.

It was also found that Mo in the slag was recovered by the soft reduction process. The MoO ₃ concentration in the slag after soft reduction is cryolite use = 0.12 wt%, NaF use = 0.12 wt%, 0.0 wt%, and about 90% is recovered from the initial concentration of 1.30 wt%. Absorbed into the Cu-Fe alloy bath.

(About hard reduction)
Looking at the concentration of P ₂ O ₅ after the hard reduction treatment, there is not much change from the initial concentration. The pig iron recovered from the copper smelting slag is likely to be low P pig iron. This is advantageous as a raw material for iron making.

On the other hand, looking at the concentration of Na 2 _O, as a concentration after treatment after a hard reduction, due to the iron component of the main component contained in the initial copper smelting slag is gradually decreases, Na ₂ O The concentration of should be higher. This is because Na ₂ O added as a flux in the slag does not change into other components and does not leak out of the system.

Here, although the results shown in Table 8 shows the behavior of the concentration of Na 2 _O gradually increases, the concentration of Na 2 _O for Tables 6 and 7 do not show the behavior as shown in Table 8.

  The purpose of this example is to recover Cu components that are valuable metals after soft reduction (reduction of numerical values in the table is an index) and to remove As components and Pb components that are harmful compounds (this is also the numerical values in the table) And the recovery of the Fe component after the hard reduction (also the decrease in the values in the table is an index).

  In Table 8, the analytical value of CuO after the oxidation treatment to the soft reduction treatment is reduced from 0.48 wt% to 0.14 wt%, and the analytical value of AsO and PbO is also from 0.33 wt% to 0.3 wt%, respectively. It can be seen that there is a decrease from 00 wt%, 0.18 wt% to 0.00 wt%.

  The analytical value of FeO after the soft reduction treatment and after the hard reduction treatment is reduced from 31.80% by weight to 3.41%.

Therefore, although Table 8 achieves the object of the present invention with high efficiency, the results of Table 6 and Table 7 are the concentrations after treatment after hard reduction when the concentration of Na ₂ O is seen. In particular, in the results shown in Table 7, the concentration of Na ₂ O after the soft reduction treatment is 8.77% by weight, which is a very small numerical value.

  And although it seems that the objective was achieved regarding the removal of harmful compounds, the concentration of FeO after hard reduction shown in Table 6 was 3.41, and the concentration of FeO after hard reduction shown in Table 7 was 6.45. Even when compared with 1.30% by weight in Table 7, the degree of recovery of the Fe component after the hard reduction is inferior.

  Furthermore, the concentrations of CuO after the oxidation treatment, soft reduction treatment, and hard reduction shown in Table 8 steadily decreased to 0.83% by weight, 0.13% by weight, and 0.00% by weight, respectively. In the experiments shown in Table 6, there are some effects such as 0.48% by weight, 0.14% by weight, and 0.01% by weight, respectively. However, in the results shown in Table 7, the Cu component remains in the slag after the hard reduction treatment, which is 0.76% by weight, 0.29% by weight, and 0.05% by weight.

This is probably due to the fact that the Na ₂ O concentration exhibits unexpected behavior at each stage as described above. More specifically, it is considered that Na ₂ O in the slag is not uniformly dispersed.

Claims (13)

A method for recovering valuable metals from copper smelting slag, comprising the following steps 1 to 3;
1 a step of oxidizing copper smelting slag with an oxygen-containing gas in the presence of flux;
2 The step of contacting the oxidized reactant obtained in Step 1 with the Cu—Fe alloy bath, and 3 The contacted Cu—Fe alloy bath obtained in Step 2 are subjected to a refining step, and valuable metals are recovered. Process.     The method of Claim 1 which further includes the process 4 which makes the copper smelting slag after the contact obtained at the said process 2 contact, and a Fe-C alloy bath.     The copper smelting slag is at least one selected from the group consisting of Cu compounds, Si compounds, Ag compounds, Au compounds, Pt compounds, Fe compounds, Mo compounds, Zn compounds, As compounds, Sb compounds, Pb compounds and Sn compounds. 3. A method according to claim 1 or 2 comprising one compound.     The said valuable metal is at least 1 compound selected from the group which consists of Fe compound, Cu compound, Ag compound, Au compound, Pt compound, Mo compound, and Sn compound in any one of Claims 1-3. the method of.     The method in any one of Claims 1-4 which implements the said process 1 on 1200 degreeC or more conditions.     The method according to claim 1, wherein the flux is a fluorine-containing compound.     The method in any one of Claims 1-6 which implements the said process 2 on the conditions of 1460 degrees C or less.     The method in any one of Claims 2-7 which implements the said process 4 on the conditions of 1150 degreeC-1550 degreeC. A method for removing harmful compounds from copper smelting slag, comprising the following step A and step B;
A: A step of oxidizing copper smelting slag with an oxygen-containing gas in the presence of a flux; and B, a step of contacting the oxidized reactant obtained in step A with a Cu—Fe alloy bath.     The method according to claim 9, further comprising a step C of subjecting the Cu—Fe alloy bath obtained in the step B to a refining step.     The method according to claim 9 or 10, wherein the harmful compound is at least one selected from the group consisting of an As compound, an Sb compound, a Zn compound, and a Pb compound.     The method according to any one of claims 9 to 11, wherein the step A is performed under a condition of 1200 ° C or higher.     The method in any one of Claims 9-12 which implements the said process B on the conditions of 1460 degrees C or less.