RU2835731C1 - Improved copper melting process - Google Patents

Abstract

FIELD: performing operations.

SUBSTANCE: invention relates to extraction of copper from secondary raw materials. Method involves melting (100) in a furnace of secondary raw material (1, 2) containing copper oxide and elemental iron to obtain concentrated intermediate copper product (3). Heat generation is ensured as a result of redox reactions converting iron into oxide and copper oxide into copper. Copper is collected in molten liquid metal phase, and iron oxides are collected in floating phase of liquid slag. At the end of melting, at least one of the liquid phases is at least partially removed from the furnace in the form of melting slag (5) and/or in the form of concentrated copper intermediate product (3). During the melting stage, excess elementary iron is maintained in the furnace in relation to the amount required for completion of redox reactions, and additional heat supply is provided due to supply of oxygen-containing gas for oxidation of excess iron.

EFFECT: stable and reliable stage of melting, direct, accurate dosing of introduced oxygen to obtain part of reaction heat, at the same time additional stage of heat transfer is not required.

42 cl, 1 dwg, 2 tbl, 1 ex

Description

FIELD OF TECHNOLOGY TO WHICH THE INVENTION RELATES

The present invention relates to the recovery of copper (Cu) together with other non-ferrous metals such as tin (Sn), lead (Pb), nickel (Ni) and zinc (Zn), mainly from secondary raw materials, by means of pyrometallurgical process steps. The present invention preferably relates to secondary raw materials, also known as recycled materials. Recycled materials may be, for example, by-products of metal producers, waste and spent materials.

PREREQUISITES FOR THE CREATION OF THE INVENTION

The present invention mainly concerns a known pyrometallurgical stage in copper production, i.e., the smelting stage. Depending on the source of the raw material, the smelting stage can be further defined as a primary or secondary smelting stage.

Smelting is a process that uses heat and chemicals, most often in a primary smelter, to a metal ore to extract the base metal. More technically, smelting is a process in which metal-containing solids are converted into liquids using a chemical reaction, resulting in metals. The process is used to extract many metals from the inert gangue component of ore, including silver, iron, copper, and other base metals. It is a form of extractive metallurgy in which chemical reactions are used to remove other elements as gases or slags, leaving behind a liquid bath containing the metal in elemental or chemically combined form, such as with sulfur, which is known as a "matte". The elemental metal is extracted in a separate phase of the molten metal. The matte also usually forms a separate liquid phase. Most ores are impure and it is often necessary to use a flux such as limestone or silica to remove the associated rocks as part of another separate liquid phase, forming a typical by-product called "slag".

In addition, when extracting copper from secondary materials, smelting can be used as a first step to extract a concentrated copper phase from secondary raw materials that may be too heavily contaminated and/or the copper content of which may be too low for direct use as feedstock for processing into anode copper. Such secondary raw materials are usually richer in copper than primary copper sources such as copper ore or even intermediate copper concentrate, which is usually first obtained from the ore, for example by flotation of the mineral raw material before the smelting step. Some of the copper in the secondary material may also already be present in elemental form and therefore not chemically bound. For these reasons, the operating conditions in the smelting step of copper from secondary raw materials differ markedly from those in the primary smelting step of copper from copper concentrate or sometimes from copper ore.

In a primary copper smelting furnace, typical feed materials are copper-bearing sulphides such as chalcopyrite (CuFeS ₂ ), bornite (Cu ₅ FeS ₄ ) and chalcocite (Cu ₂ S). Their reaction with oxygen (oxidation) removes sulphur as SO ₂ in the off-gas and forms a "matte" phase (Cu ₂ S.FeS) together with a slag phase (FeO.SiO ₂ ) which is formed by reaction with the addition of silica. In a second stage, usually also in the smelting furnace, the FeS is removed by further reaction with oxygen and silica to form more slag and more SO ₂ gas, leaving the so-called "white metal" (Cu ₂ S), usually containing less than 1% remaining Fe. The latter is then oxidised with an oxygen-containing gas, preferably air, according to the reactions

Cu ₂ S+O ₂ → 2Cu+SO ₂ ,

Cu ₂ S+3/2 O ₂ → Cu ₂ O+SO ₂

Cu ₂ S+2 Cu ₂ O → 6 Cu+SO ₂

This primary copper smelting process is usually carried out in a so-called Pierce-Smith converter, producing so-called blister copper, which contains nickel and precious metals, a slag containing most of the Fe, Zn together with 2-15% Cu, furnace dust (including most of the Sb, As, Bi, Cd, Pb) and off-gases (again containing SO ₂ ). The large amount of sulfur that is released in these process stages is removed as SO ₂ and recovered as sulfuric acid. Thus, the primary stage of copper smelting is usually and primarily a strong oxidation stage. Primary copper smelting is described, for example, in patent document JPS61531 (A) or its patent version JPH0515769 (B2). The treatment of the copper matte phase is described, for example, in patent documents CN101871050 A and GB2462481 A. Patent document US 3954448 describes a process for further treatment of copper matte or slag from the primary stage of copper smelting.

Patent document JP 2003253349 describes a primary copper smelting process in which the matte feedstock also contains iron sulphide. In the first stage, this iron sulphide is selectively oxidised to iron oxide using oxygen-enriched air. The iron oxide and added silica form a separate slag phase which is removed from the furnace prior to further processing of the copper sulphide in the second stage. The amount of iron sulphide available and its reaction with FeO may not be sufficient to produce the reaction heat required to maintain the furnace temperature during this first stage, particularly when cold feedstock is to be processed. In the first stage, additional metallic iron is added per ton of matte processed to counteract further oxidation of the iron to magnetite (Fe ₂ O ₃ ), which would otherwise increase the viscosity of the slag and impair subsequent phase separation and slag removal at the end of the first stage. Under oxidizing conditions, which are usually dominant at the primary copper smelting stage, additional metallic iron is oxidized and this reaction releases additional heat.

The metals in typical feedstocks for the secondary copper smelting stage are mainly present as oxides, although small amounts of sulphides may be present. Thus, an important difference from the primary copper smelting stage is the absence of any copper matte as an intermediate product. Some copper in the feedstock may in fact already be present in its elemental form, but in too low a concentration or in a form less suitable for pyrometallurgical copper refining and even less so for hydrometallurgical extraction (leaching + electrowinning). The copper in the oxides is then reduced in the smelting stage by the addition of a reducing agent, such as a carbon source such as coke and/or metallic iron, usually in the form of scrap iron. Thus, another important difference from the primary copper smelting stage is the reducing conditions under which the secondary smelting stage is carried out.

Patent document US 3,682,623 (Ludo Dierckx et al.) describes a process for refining copper starting from secondary raw materials, in which the first stage is a smelting stage, i.e. a reduction stage, carried out in a smelting furnace in which copper-containing materials are heated together with a solid material containing metallic iron under an oxygen-rich neutral flame, and with gentle stirring a bath is obtained from this charge, including a slag phase. A small amount of alkaline or neutral flux may be added in order to optimize the density and viscosity of the slag formed. Additional silica may also be added to absorb iron compounds formed in the reduction reactions. It has been found that when the temperature of the smelting bath in the furnace is increased, chemically bound copper, lead, tin or nickel in the charge are reduced by metallic iron in the solid state, forming a molten metal called "black copper" and a molten slag containing iron silicate. Typical reactions of this method are said to include:

MeO+Fe → FeO+Me

(MeO) _x SiO ₂ +x Fe → (FeO) _x SiO ₂ +x Me

x FeO+SiO ₂ → (FeO) _x SiO ₂

These reactions confirm that the so-called "melting" stage in US Patent No. 3,682,623 qualifies as a "smelting" stage in the context of the present document. These reactions are exothermic, and it is stated that the heat of reaction rapidly increases the temperature of the charge. Once the material has melted to such an extent that it flows easily along the wall of the vessel, the stirring in the vessel can be increased. At the end of the reduction stage, black copper and smelting dross are formed, which can be separated from each other by gravity and can be removed from the furnace separately.

Throughout the reduction operation the temperature is maintained as low as practicable while maintaining the slag in a liquid state. The fuel feed should be controlled so that the temperature of the reaction mass does not exceed about 1300°C for any significant period during the furnace cycle. It is preferable to maintain the temperature not much above the temperature at which the slag becomes substantially liquid. A bath temperature of about 1180°C is stated to be satisfactory for ordinary charge materials, but lower temperatures may be used if borax is used as a flux.

The low temperature not only minimizes the evaporation of lead and tin, but also limits the dissolution of solid iron in the resulting copper melt. It is argued that the presence of a significant amount of solid iron is essential to ensure rapid and complete reduction of the slag. The dissolution of iron must also be minimized to maintain high solubility of lead and tin in the resulting black copper. As the reduction reactions proceed and the solid iron-containing material gradually dissolves in the molten metal, additional solid material containing metallic iron may be advantageously added after the completion of the smelting to effect the final reduction of the copper, tin, lead and zinc remaining in the slag. In general, the excess iron that remains in the furnace is utilized and at least some of it is dissolved in the black copper. Zinc is volatilized from the furnace, but a significant amount of zinc also remains in the black copper at the end of the smelting stage.

At the end of the so-called smelting stage, when the final reduction stage was completed, as determined by further slag analysis, the smelting slag was poured out of the furnace from above and granulated. After the smelting slag had been poured out of the furnace, the resulting black copper was subsequently pre-refined in the same furnace with the addition of additional secondary raw materials that were already quite rich in copper, and a strongly oxidizing flame was used for this pre-refining. This pre-refining stage is therefore no longer part of the preceding smelting stage, which is a reduction stage characterized by a reducing environment.

In Example 1 of US 3,682,623, the furnace charge is melted "under a neutral, oxygen-rich flame" (Col. 15, lines 1-2), which refers to a neutral flame using oxygen-rich air. After adding additional copper-iron scrap, the slag was further reduced in a slightly reducing flame (Col. 15, lines 33-35). Most of the zinc present was vaporized and collected as dust in the exhaust system. The slag was then poured out of the furnace from the top and granulated.

Patent document DE 10 2012 005 401 A1 describes a bath smelting furnace in which a copper-containing substance, preferably a secondary raw material containing copper, is subjected to a smelting process that operates on oil and/or gas together with air and/or oxygen, which are introduced into the bath by means of a submerged injection lance. In the smelting step, a primary slag is formed, which contains comparatively few impurities and is removed from the process, as well as a second slag for further processing, which is transferred from the bath smelting furnace to a rotary drum furnace. The rotary drum furnace is provided with a burner at one end, into which oil or gas can be fed, and optionally oxygen from an oxygen store. Further processing occurs in stages and successively produces anode copper, black copper, a crude tin mixture, which can be further treated with silicon, and a final slag. Coal is introduced into the rotary drum kiln at each stage of the process. At each stage of the process, in accordance with DE 10 2012 005 401 A1, the kiln is heated by burning the fuel with air and/or oxygen.

Patent document EP 0185004 describes a process in which an oxidative smelting step, carried out on secondary materials in an attempt to increase the yield of valuable metals, results in a liquid bath from which tin and zinc are subsequently removed by evaporation in two stages, after which the iron-containing slag is drained off and the copper-containing metallic phase is retained. This metallic phase is further processed in order to first separate the lead-silicate slag, the low-nickel blister copper and the copper-nickel oxide bath, which can then be reduced to form a copper-nickel alloy.

Patent document US 2017/0198371 A1 addresses the possible large fluctuations in the organic content of the feedstock for the smelting furnace and their impact on the productivity of the process. This document proposes to remove the organic components in a batch mode in the first stage, thereby already producing so-called “black copper”, which can be converted into blister copper in a subsequent stage by further oxidation. Another product of the first stage is a final slag with a low metal content. The document states that “an appropriate amount of oxygen is blown into the working chamber”. In this process, “the composition of the slag and the content of valuable metals still present in it are monitored during the smelting process by taking samples and quickly analyzing them”.

A disadvantage of the melting stage according to US 3682623 is that for most of it, a significant part of the heat input is provided by a neutral flame operating on oxygen-enriched air. This requires a large source of fuel, air and pure oxygen, which is associated with complexity, additional equipment and operating costs.

The heat input from the flame above the furnace to the melting bath leaves much to be desired because the heat must be transferred from the flue gases to the liquid bath. The heat transfer from the gas to the liquid is quite slow and the contact surface area between the burner at the top of the furnace and the liquid bath remains limited. With a submerged burner, gravity causes the gas to rise quickly and leave the liquid phase. Thus, there is a relatively short contact time between the combustion products and the melting bath. Thus, the heat input from the flame to the bath occurs mainly by radiation. Consequently, the heat input by the flame is not very efficient and most of the potential heat input from the flame escapes from the furnace into the exhaust gases, where it represents an additional load on the exhaust gas cooling system.

The large volumes of exhaust gases created by the flame also require the installation and operation of a large exhaust gas cleaning system.

The inventors have discovered that there remains a need for a more convenient and more efficient introduction of heat into a secondary copper smelting furnace while maintaining or even improving the means of temperature control during the smelting stage.

The present invention aims to eliminate or at least alleviate the above-described problem and/or to provide improvements.

ESSENCE OF THE INVENTION

According to the present invention, there is provided a process defined in the appended claims.

In one embodiment, the present invention provides a method for recovering copper from secondary feed materials comprising the step of, in at least one batch of feed, smelting a feed comprising the feed materials in a furnace to recover a concentrated copper intermediate product therefrom,

in which a feedstock of starting materials is gradually introduced into the furnace, the feedstock comprising copper and optionally at least one metal which, under the operating conditions of the furnace, is more noble than tin, at least partly in the form of an oxide,

wherein the raw material further comprises iron and optionally at least one metal or compound which, under furnace conditions, is at most as noble as iron or zinc, wherein the iron and this metal, at most as noble as iron or zinc, are at least partially present in elemental form,

in which heat is generated inside the furnace by oxidation-reduction reactions converting elemental iron and metals or compounds at most as noble as iron or zinc into oxides, and converting oxides of copper and metals more noble than tin into the elemental metal,

in which the elemental metals are at least partially collected in the molten liquid metal phase and the oxides are at least partially collected in the floating liquid slag phase,

in which the liquid phases can be separated and at the end of the smelting stage at least one of the liquid phases is at least partially removed from the furnace in the form of smelting slag and/or in the form of a concentrated copper intermediate product,

characterized in that

during the smelting stage, the furnace is maintained in excess of the elemental form of iron and of metals or compounds which under the furnace conditions are at most as noble as iron or zinc, relative to the amount required to complete the oxidation-reduction reactions, and

additional heat supply to the furnace is provided during the melting stage by feeding oxygen-containing gas to oxidize excess iron and metals or compounds, at most as noble as iron or zinc, present in the furnace and optionally to burn a combustible source of carbon and/or hydrogen that may be additionally introduced into the furnace.

It is preferable to maintain an excess of iron and optionally metals or compounds which under furnace conditions are at most as noble as iron or zinc by adding to the furnace as part of the feedstock at least one additional starting material rich in iron and/or at least one metal or suitable compound.

The inventors have found that maintaining in the furnace an excess of iron and optionally other metals and/or compounds which are at most as noble as iron or zinc provides a very convenient way of regulating the heat input and, consequently, the temperature inside the smelting furnace, i.e. by controlling the input of oxygen-containing gas into the furnace, because the oxygen in this gas becomes available for the oxidation of the excess of elemental iron and/or other metals or compounds which are at most as noble as iron or zinc. The inventors have found that this method provides a direct, accurate and correct dosing of the introduced oxygen for obtaining a part of the reaction heat and, what is more advantageous, that part which becomes easily, directly and completely available in the liquid bath in the furnace at the level where this introduced oxygen is most desirable for obtaining a part of the reaction heat and, what is more advantageous, i.e. at the interface between the metallic phase and the slag phase, where the oxidation-reduction reactions and phase exchanges are expected.

The heat from the oxidation of iron and other metals and/or compounds, at most as noble as iron or zinc, by reaction with oxygen from the oxygen-containing gas is generated in the bath itself and does not require any additional heat transfer stage. This heat of reaction is completely and immediately dissipated in the melting bath.

The inventors have found that the temperature control in the smelting furnace is simple and very sensitive thanks to the present invention. This is very advantageous because when the temperature in the molten bath increases, more iron dissolves in the molten liquid metal and becomes available for oxidation by the available oxygen, which in the case of abundant oxygen would release even more heat and could sharply increase the temperature.

The present invention is able to avoid this risk of temperature going out of control, since in the process according to the present invention the supply of oxygen as part of the oxygen-containing gas is strictly controlled. If the feedstock at a certain point introduces a higher amount of oxygen, which becomes available to participate in oxidation-reduction reactions, and if the heat released by these additional oxidation-reduction reactions leads to an increase in the temperature of the smelting bath, the temperature of the smelting bath can be easily brought under control by reducing the injection rate of the oxygen-containing gas, and any risk of temperature going out of control can be easily avoided or at least significantly reduced.

Another advantage of the present invention is that the oxidation of iron and other metals or compounds that are at most as noble as iron or zinc does not typically result in the formation of a large amount of exhaust gas, unlike the combustion of natural gas or other carbon and/or hydrogen-containing fuel, and that the furnace exhaust gas cleaning system associated with the smelting equipment in which the process according to the present invention is carried out can be designed to be smaller in size and therefore requires lower capital investment and also requires lower operating costs during operation. Another advantage of the smaller volume of exhaust gases is that the less valuable tin, lead and zinc are usually evaporated and therefore do not need to be captured in the exhaust gas cleaning system.

Thus, the inventors have found that the release of heat by oxidation, for example, the conversion of excess iron to iron oxide by introducing gaseous oxygen into the smelting bath, is much more efficient than the flame combustion of a combustible carbon and/or hydrogen source in a furnace. The inventors have determined that about 80% of the introduced oxygen reacts with compounds in the molten bath of the furnace, and that the heat released by these reactions remains in the molten bath, which is a very high yield compared to the heat that can be caused by the combustion of a hydrocarbon fuel such as natural gas, even if this combustion occurs on pure oxygen or oxygen-enriched air. The inventors believe that this difference is due to the fact that the conversion of iron to iron oxide occurs in the molten bath itself, while the combustion of natural gas occurs in the gas phase, and the heat of combustion must still be transferred to the liquid phase in order to contribute to the enthalpy content of the molten bath. In addition, such combustion may not necessarily be complete.

The inventors have also found that suitable sources of elemental iron and metals or compounds, at most as noble as iron or zinc, are readily available from a large number of sources and can be readily obtained under economic conditions that make this method of heat supply more economically advantageous than heat supply using a neutral flame, even based on operating costs alone. In addition, the carbon footprint of the method according to the present invention is lower than that of the above-described prior art method.

Although it is within the scope of the present invention, the inventors prefer not to use the option of providing a portion of the total heat input to the furnace by burning a combustible carbon and/or hydrogen source in the furnace. The inventors have found that under certain economic conditions, using this option may be advantageous, but the inventors always, including when using this option, prefer to control the furnace temperature by supplying oxygen-containing gas due to its higher level of convenience, easier control and less risk of temperature getting out of control.

Another advantage of the present invention, associated with the excessive presence of iron dissolved in the layer of molten liquid metal inside the furnace, is that around the submerged tuyeres through which oxygen-containing gas can be introduced, a layer of solid iron and/or iron oxide is formed, which forms an additional protection of these tuyeres from wear, since these tuyeres are cooled by a gas flow, usually colder than the smelting bath. This protective layer is usually in the form of a hollow mushroom and is formed due to the fact that the tuyeres themselves are colder, and the molten liquid metal around the tuyeres becomes colder, as a result of which the solubility of iron in the concentrated molten liquid copper phase decreases, iron precipitates and adheres to the outer surfaces of the tuyeres, except for the throat through which the gas is blown.

High sensitivity of temperature control is advantageous because if the temperature of the molten liquid increases, this protective layer can dissolve again and the tuyere can lose its protective layer, which can lead to serious damage and production losses. High sensitivity of the temperature control system means that the risk of such damage to tuyeres and the associated production losses can be significantly reduced and preferably prevented.

The inventors have found that the beneficial effects of the present invention result in a more stable and reliable smelting stage. The smelting stage is usually the very first stage of a more complex overall pyrometallurgical process.

The complete process may, for example, process the products of the smelting stage in accordance with the method according to the main claim of the invention into their derivatives.

The slag from the smelting stage may preferably be further processed, such as by evaporation, to produce a slag that not only poses fewer disposal concerns and/or is used for more valuable end uses, as described later in this document.

The concentrated intermediate copper product, preferably after separation from the smelter slag formed in the smelting stage, may be further processed, such as by refining to produce a more concentrated refined copper product that becomes suitable for higher value end uses, optionally by casting copper anodes as feedstock for electrolysis, which may ultimately result in high purity copper cathodes meeting many if not all current industry standards for more demanding copper end uses.

Further processing of the concentrated intermediate copper product and/or slag from the smelting stage may result in the production of other valuable by-products from the refined copper product.

Such a valuable by-product may be, for example, a crude solder stream that may be obtained from copper refining slag that may be formed as a result of refining a concentrated intermediate copper product. Such crude solder may be further purified and/or adjusted, i.e. purified by removing elements that may affect subsequent processing and/or impair or hinder specific applications of the final products obtained from the solder stream. These valuable by-products may include, for example, at least one of the products belonging to the list consisting of a soft lead base product, a hard lead base product, silver-rich anode slime and a high-grade tin base product, as will be explained further in this document.

The inventors believe that the beneficial effects provided by the present invention are transferred to all further stages, up to obtaining the derivative products of the melting stage, which were listed above. The increased stability and reliability of the melting stage provides the advantage that the subsequent processes in which these derivatives are produced provide a more stable and reliable flow of raw materials originating from the melting stage, which makes their own operation more stable and reliable. This allows for the production of final products of more stable and reliable quality. In addition, this allows for a reduction in the burden on monitoring the method and/or the attention of the operator, and also increases the possibility of electronic monitoring and control of each stage in these methods, as well as the method as a whole.

BRIEF DESCRIPTION OF DRAWINGS

Fig. 1 shows a flow chart of a method incorporating the process according to the present invention as part of a general method for extracting non-ferrous metals from secondary raw materials.

DETAILED DESCRIPTION

The present invention will be further described in the form of specific embodiments with possible references to specific drawings, but the present invention is not limited thereto, but is defined only by the claims. The drawings described are schematic and are not limiting. In the drawings, the size of some of the elements may be exaggerated and not shown to scale for illustrative purposes. The dimensions and relative sizes in them do not necessarily correspond to the actual ones.

In addition, the terms "first," "second," "third," and the like in the description and claims are used to distinguish between like elements and not necessarily to describe a sequential or chronological order. These terms are interchangeable under appropriate circumstances, and embodiments of the present invention may operate in other sequences than those described and/or illustrated herein.

In addition, the terms "top," "bottom," "above," "below," and the like in the description and claims are used for descriptive purposes and not necessarily to describe relative positions. These terms are interchangeable under appropriate circumstances, and the embodiments of the present invention described herein may operate in orientations other than those described or illustrated herein.

The term "comprising" as used in the claims shall not be construed as being limited to the elements that are recited in context with it. It does not exclude the presence of other elements or steps. Thus, it shall be interpreted as specifying the presence of the claimed features, integers, steps or components as mentioned, but does not preclude the presence or addition of one or more other features, integers, steps or components, or groups thereof. Thus, the scope of "an article comprising means A and B" cannot be limited to an object that consists exclusively of means A and B. It only means that A and B are the only elements associated with the subject matter of the present invention. Accordingly, the terms "comprising" and "including" embrace the more restrictive terms "consisting essentially of" and "consisting of". Therefore, by replacing the phrases "comprising" and "including" with "consisting of," these terms form the basis of preferred but limited embodiments that are also offered as part of the content of this document regarding the present invention.

Unless otherwise expressly stated, all ranges given herein include the end points of the stated ranges of content values of the components of the composition, expressed as wt.% of each ingredient.

As used herein, the terms "mass percent", "wt%", "percent by weight", "parts per million by weight", "parts per million by weight" or "ppm" and variations thereof refer to the concentration of a substance as the mass of that substance divided by the total mass of the composition and multiplied by 100 or 1,000,000, as the case may be, unless explicitly stated otherwise. As used herein, "percent", "%" is intended to be synonymous with "mass percent", "wt%", etc.

It should be noted that the grammatical forms of the singular used in this document and in the appended claims also include the plural, unless the context clearly indicates otherwise. Thus, for example, a reference to a composition comprising a "compound" includes a composition comprising two or more compounds. It should also be noted that the term "or" is generally used in the meaning of "and/or", unless the context clearly indicates otherwise.

In addition, each compound used herein may be referred to interchangeably by its chemical formula, chemical name, abbreviation, etc.

Metals and compounds, at most as noble as iron or zinc, are compounds which, under furnace conditions, have at least the same or even a higher affinity for oxygen than iron or zinc, and therefore also copper, nickel, tin and lead. This definition refers to "iron or zinc" because the relative positions of zinc and iron in their affinity for oxygen under furnace conditions are very close and can even change depending on furnace conditions. In order to be correct and exhaustive in this definition, it is necessary to refer to both of these metals. Under furnace conditions, these selected metals or compounds enter quite easily into oxidation-reduction reactions on the oxidizing side, which is part of the present invention. Suitable metals are, for example, elemental zinc and iron themselves, aluminum, silicon and calcium. Suitable compounds may be, for example, metal silicides, preferably metal silicides that are already suitable in themselves, such as iron silicide (FeSi), but bimetallic or polymetallic compounds may also be suitable, including mixtures, such as SnAl, CuFe, FeSn, or alloys, such as brass (ZnCu). Other suitable compounds may be metal sulfides, such as FeS, ZnS and/or sulfides of other metals that are at most as noble as iron or zinc.

Metals and compounds more noble than copper are compounds which, under furnace conditions, have a lower affinity for oxygen than copper. These starting materials enter into oxidation-reduction reactions on the reducing side as part of the present invention quite easily and lead to the release of the corresponding metal in its elemental form. Suitable are, for example, silver, gold, other precious metals, including metals of the platinum group, their alloys and mixtures, including those containing other metals.

In the context of the present invention, the terms "smelter", "melt", "melting" or similar derivatives of "melting" mean a process that involves much more than simply changing the state of a compound from solid to liquid. In the pyrometallurgical smelting stage, several chemical processes occur that transform certain chemical compounds into other chemical compounds. Important of these transformations may be oxidations, possibly associated with the formation of an oxide, or reductions, in which the oxidation state of certain atoms changes. In this document, the term "smelter" means the furnace in which this stage of the process occurs.

In the context of the present invention, the terms "dross" or "drosses" mean a substance, often pasty, that is formed as a result of a working stage and that separates from another liquid phase, usually under the action of gravity, and usually floats to the surface. The dross or drosses can usually be skimmed or removed from the liquid that is underneath them.

The term "solder" in the context of the present invention means a metal composition that is rich in tin and/or lead, but which may also contain other metals. The solder is characterized by a relatively low melting point, which makes this composition suitable, after heating it to a relatively limited temperature, for forming a metallic connection between two other metal parts upon cooling, the so-called "soldering".

In this document, unless explicitly stated otherwise, the amounts of metals and oxides are expressed in accordance with common practice in pyrometallurgy. The presence of each metal is usually expressed in its total presence, regardless of whether the metal is present in its elemental form (oxidation state = 0) or in any chemically combined form, usually in oxidized form (oxidation state > 0). For metals that can be relatively easily reduced to their elemental forms, and which can occur as molten metal in a pyrometallurgical process, it is quite common to express their presence in terms of their elemental metallic form, even when describing the composition of a slag in which most of such metals may actually be present in oxidized form. Accordingly, the composition of a slag, such as the slag obtained in the process according to the present invention, determines the content of Fe, Zn, Pb, Cu, Sb, Bi as elemental metals. Less noble metals are more difficult to reduce under non-ferrous pyrometallurgical conditions and they occur mainly in oxidized form. These metals are usually expressed as their most common oxide form. Therefore, the slag composition usually gives the content of Si, Ca, Al and Na, respectively expressed as SiO ₂ , CaO, Al ₂ O ₃ , Na ₂ O.

Metallurgical slag is usually not a pure substance but a mixture of many different components. Consequently, metallurgical slag does not have a clear melting point. In this field, it has become common to use the term "liquidus temperature", which is the temperature at which the slag becomes completely liquid.

In one embodiment of the method according to the present invention, the feedstock further comprises at least one second metal selected from the group consisting of nickel, tin and lead. Feedstocks containing at least one second metal selected from this list are of great interest for the recovery of copper therefrom, but the presence of the second metal may create additional problems or difficulties compared to feedstocks that do not contain the second metal. The inventors have found that the smelting step is a very suitable process step for introducing feedstocks containing at least one of such second metals. The inventors have found that this at least one second metal may also be present in the feedstock as its oxide or in another form capable of participating in oxidation-reduction reactions under furnace conditions and releasing the metal in its elemental form. The inventors prefer the oxide form in this context because of its higher availability under favorable conditions, because of the reaction heat that is released by the oxidation-reduction reactions in which it participates, and also because of its contribution of oxygen to the furnace, which reduces the amount of oxygen that must be introduced. The reduction in the need for oxygen supply also reduces the gas consumption in the melting stage, which is advantageous since the reaction rate is not limited by the rate of external supply of gaseous oxygen, but is limited only by the kinetics of the reaction. Less gas input can also mean less exhaust gases from the furnace that need to be cleaned, as well as less entrainment of solid particles with them.

In one embodiment of the method according to the present invention, in which the feedstock contains at least one second metal, the concentrated intermediate copper compound further contains at least one second metal. The inventors have found that the smelting step can be carried out in such a way that a large part of the second metal is also recovered as part of the concentrated intermediate copper compound, most easily by, for example, bringing the oxidation-reduction reactions in the smelting furnace to an appropriate degree. The inventors have found that this feature offers the advantage that also at least one second metal can be additionally recovered as part of the primary product of the desired high quality.

In one embodiment of the method according to the present invention, the feedstock comprises scrap iron, silicon, zinc and/or aluminum, more preferably scrap iron. The inventors have found that this scrap can be easily dosed with sufficient accuracy by mixing appropriate amounts with other feedstocks in a batch feed. The inventors can also add this scrap as an additional feedstock stream to the furnace. The inventors have found that scrap materials such as scrap iron and scrap aluminum, but to some extent also scrap silicon, can be easily obtained in suitable amounts and under economically favorable conditions. The inventors have also found that a separate addition of scrap as an additional feedstock stream to the furnace, preferably scrap iron, offers the advantage of being able to very conveniently control and maintain an excess of elemental iron and/or metals and/or compounds that are at most as noble as iron or zinc under the furnace conditions.

In one embodiment of the method according to the present invention, the process further comprises a step of removing smelting slag from the furnace, at least partially. The inventors prefer to remove at least a portion of the smelting slag from the furnace before the next charge is started. If the raw material that is available during the new feed charge contains a significant portion of the fine fraction, the inventors prefer to maintain a slag layer in the furnace, because this layer provides a suitable cover under which the fine part of the raw material or the raw material containing a significant portion of the fine fraction can be introduced without creating an excessive risk that the fine particles of the raw material will be carried away by the furnace exhaust gases and create an additional load and/or interference for the exhaust gas treatment system. If the available raw material contains a significant portion of the coarse fraction, the inventors prefer to remove substantially all of the formed slag from the furnace before the next charge is started. This has the advantage that a larger volume of the furnace can be available for the next batch of charges, which has a beneficial effect on the throughput and/or productivity of the smelting furnace. The inventors have discovered that the step of removing slag from the furnace can be carried out several times during the same furnace charge.

In one embodiment of the method according to the present invention, the method comprises the step of removing at least a part of the concentrated intermediate copper product from the furnace, preferably only a part. The inventors prefer to ensure that there is a suitable physical presence of molten metal in the furnace at the start of the next batch of feed or a cycle comprising a series of feed charges. This molten metal is then readily available as a hot liquid already at the start of a new batch of feed or the entire cycle for receiving and wetting the solid feedstock, as well as possibly additional quantities of elemental iron and metals or compounds that are at most as noble as iron or zinc under the operating conditions of the furnace, which may be desirable or necessary for easily creating and/or storing an excess of these additives for the method according to the present invention. The iron dissolved in this molten metal is readily available for reaction with the oxygen that is introduced into the liquid bath and thus for the immediate release of the reaction heat. Another advantage is that the solid iron that can be added to the furnace at the start of a new feed charge remains floating in the molten metal phase, exactly where it can fully contribute to the oxidation-reduction reactions that are provided for in the process. The inventors have found that retaining a portion of the concentrated intermediate copper product in the furnace when starting a new batch of feed for the furnace significantly reduces the delay time before the furnace can again operate at high productivity, and therefore provides a significant improvement in the productivity of the smelting stage. The inventors prefer to remove a portion of the molten metal formed during the previous feed charge before starting the next charge. The inventors have found that the step of removing a portion of the concentrated intermediate copper product from the furnace can even be carried out several times during the same furnace charge.

The inventors prefer to carry out the smelting step as nearly as possible in a semi-continuous manner, whereby suitable material can continue to be added to the furnace until the useful volume of the furnace is completely used. When the slag and metal phases have reached the desired quality, at least a major part of the slag can first be removed from the furnace, for example by pouring through a feed opening when tilting the furnace, and then a significant part of the liquid molten metal phase can also be removed in the same way, if all the slag has been removed, or drained through a bottom taphole suitably located in the wall of the furnace. For the reasons explained above, a suitable part of the molten metal preferably remains in the furnace when the introduction of the next batch of charge into the smelting furnace is started. The inventors have found that this operation can be continued for a very long period of time, and can only be terminated or stopped for external reasons, or when maintenance of the smelting furnace is necessary. The inventors have found that this operation can be further improved by preparing suitable pre-mixed batches of feedstock in terms of the composition and size of the solids in the batch. The inventors have found that this can provide the advantage of much more stable operation in terms of time in the sequence of stages, as well as the quality of the concentrated copper intermediate product that is removed from the furnace each time as the main product.

In one embodiment of the method according to the present invention, the iron and compounds at most as noble as iron or zinc, which are introduced together with the raw material, comprise solid iron, solid silicon, solid zinc and/or solid aluminum, preferably comprise copper/iron scrap, silicon-containing scrap, zinc-containing scrap, and/or aluminum-containing scrap. The inventors have found that these sources of iron, silicon, zinc and aluminum are readily available from various sources. In addition, they may contain small amounts of other metals that can be and are recoverable in their elemental form during and after the smelting stage. Such other metals may include tin, lead and nickel. They may also include traces of more noble and even precious metals (PM) such as silver or gold, and even platinum group metals (PGM) such as ruthenium, rhodium, osmium, palladium, iridium and platinum itself, even very small quantities of which are worth extracting due to their rarity and high economic value.

In one embodiment of the method according to the present invention, the raw material is at least partially solid, and therefore the solid raw material is fed into the furnace gradually, preferably continuously, preferably during most of the charging of one batch of the charge, and more preferably during most of the entire melting cycle, preferably by means of at least one belt conveyor and/or a vibrating conveyor. As explained elsewhere in this document, the inventors prefer to feed at an early stage of the furnace charging and/or the melting cycle a coarse part of the available raw material until a sufficiently thick layer of metallurgical slag has formed as a cover layer on the molten metal phase in the furnace. If this slag layer becomes accessible from the very beginning of the feedstock feed or after it has been formed during the operation of the furnace on the coarse-grained portion of the accessible feedstock, the inventors prefer to also introduce the fine-grained portion of the accessible feedstock into the furnace pneumatically through a tuyere which is immersed in the liquid bath and releases the fine-grained material approximately at the boundary between the molten metal phase and the upper molten slag phase, since this offers the advantage of a low risk of entrainment of fine feedstock particles in the exhaust gases.

In one embodiment of the method according to the present invention, the feedstock feed rate is maintained below the rate at which the heat release becomes insufficient to melt the solid feedstock and/or to bring the feedstock to the desired furnace temperature. The inventors prefer to avoid, as far as possible, the risk that the enthalpy balance of the furnace becomes insufficient, since the heat release is insufficient to heat and melt the feedstock, in which case there is a risk of a drop in the furnace temperature. The inventors have found that it is advantageous to control the feedstock feed rate, and for example iron can be added at a sufficient rate so that the excess of elemental iron and metals and compounds that are at most as noble as iron or zinc under furnace conditions remains high enough, combined with a sufficient amount of oxygen to produce sufficient reaction heat to achieve rapid heating and melting of the feedstock.

In one embodiment of the method according to the present invention, at least a portion of the feedstock is in the form of a finely divided portion, which has an average particle size of no more than 10 mm, preferably no more than 3.36 mm. The inventors have found that finely divided feedstock materials containing copper and other metals of interest in the method according to the present invention are generally difficult to process by alternative methods and therefore can be found in significant quantities at economically attractive conditions. The inventors have found that such materials can be easily processed by the method according to the present invention. The inventors prefer to introduce such finely divided feedstock portions into the furnace only after a continuous layer of molten slag has appeared in the furnace, floating on top of the molten metal phase. The inventors prefer to introduce a finely divided portion of the raw material approximately at the interface between the molten metal and the molten slag so that the slag layer can act as a protective layer capable of capturing any fine particles before they can reach the furnace gas phase and be at risk of being carried away by the exhaust gases and not participating in the process inside the furnace.

In one embodiment of the method according to the present invention, wherein the feedstock comprises a finely ground portion of the feedstock, the material of the finely ground portion of the feedstock is conveyed pneumatically and introduced into the furnace. The inventors have found that this is a very convenient way of introducing such a finely ground portion of the feedstock, and this method makes it possible to introduce this portion at the most advantageous location, i.e., immediately above the interface between the liquid metal and the liquid slag, where any elemental iron present, such as scrap iron, also typically floats and where most of the chemical reactions take place.

In one embodiment of the method according to the present invention, wherein the feed comprises a finely divided feedstock portion, the finely divided feedstock portion material is introduced into the liquid slag phase and above the metal phase of the liquid bath. The inventors have found that the slag phase is located just above the interface between the liquid metal and the liquid slag, where any elemental iron present, such as scrap iron, also typically floats. This is where most of the chemical reactions occur, and where most of the reaction heat is released.

In one embodiment of the method according to the present invention, in which the feed comprises a finely ground portion of the raw material, the average composition of the finely ground portion of the raw material fed to the entire batch of feed for melting in the furnace meets at least one, and preferably all, of the following conditions after heating to 1150°C:

- contains at least 5% by weight of the total amount of metal, preferably at least 5% by weight of the total amount of copper, nickel, tin, lead and zinc, preferably at least 6% by weight, more preferably at least 7% by weight, even more preferably at least 8% by weight, preferably at least 9% by weight, more preferably at least 10% by weight of the total amount of metal, preferably of the total amount of copper, nickel, tin, lead and zinc,

- contains at most 70.0 wt.% copper (Cu), preferably at most 65.0 wt.%, more preferably at most 60.0 wt.%, even more preferably at most 55.0 wt.%, even more preferably at most 50.0 wt.%, preferably at most 48.0 wt.% copper, and optionally at least 10 wt.% copper, preferably at least 15 wt.%, more preferably at least 20 wt.%, even more preferably at least 25 wt.%, even more preferably at least 30 wt.%, preferably at least 35 wt.%, more preferably at least 40 wt.% and even more preferably at least 42.0 wt.% copper,

- contains no more than 2.00 wt.% nickel (Ni), preferably no more than 1.50 wt.%, and more preferably no more than 1.00 wt.% nickel,

- contains at least 0.50 wt.% and no more than 10.00 wt.% lead (Pb), preferably at least 1.00 wt.%, more preferably at least 1.50 wt.%, and optionally no more than 9.00 wt.%, preferably no more than 8.00 wt.% lead,

- contains no more than 15.00 wt.% tin (Sn), preferably no more than 14.00 wt.%, more preferably no more than 13.00 wt.%, and even more preferably no more than 12.00 wt.% tin,

- contains no more than 2.00 wt.% antimony (Sb), preferably no more than 1.50 wt.%, and more preferably no more than 1.00 wt.% antimony,

- contains no more than 7.0 wt.% iron (Fe), preferably no more than 6.0 wt.%, more preferably no more than 5.0 wt.%, even more preferably no more than 4.0 wt.%, and even more preferably no more than 3.50 wt.% iron, and

- contains no more than 55 wt.% zinc (Zn), preferably no more than 50 wt.%, more preferably no more than 45 wt.%, even more preferably no more than 43 wt.%, even more preferably no more than 40 wt.%, and even more preferably no more than 35.0 wt.% zinc.

The inventors have found that said finely ground portion of the feedstock is very suitable for the process according to the invention due to the presence of metals that are of interest for recovery as high quality products after the process and/or due to the presence of metals that can release heat of reaction as part of the process according to the invention, while the finely ground portion of the feedstock has a sufficiently low content of these metals so that this portion does not represent sufficient economic interest for alternative processes for the recovery of metals from primary and/or secondary raw materials, and therefore this feedstock can be found at economically attractive conditions, which provides a significant benefit when processed by the process according to the invention. Compliance with the upper limit of the sulfur content within one of the conditions in the above list further allows to avoid the formation of a separate copper matte phase, and thus clearly distinguishes the process according to the invention, including this feature, from copper smelting processes in which the matte phase is formed as one of the products or intermediate products.

In one embodiment of the method according to the present invention, the feedstock comprises at least one recycled material from the processing of the molten liquid metal phase and/or the liquid slag phase formed by this method. The inventors have found that the smelting step is a very convenient step for the return of by-products that may be formed as a result of further processing of the molten liquid metal phase and/or the liquid slag phase formed in the smelting step. Such further processing may take place after the smelting step in the same furnace, but preferably after the smelting step and in other equipment. Examples of such further processing are further explained below in this document.

In one embodiment of the method according to the present invention, wherein the feedstock comprises at least one recycled material, the at least one recycled material comprises at least one material selected from rejected anodes or other products containing copper, tin and/or lead, an oxide or sulfide of a metal, preferably copper, nickel, tin, lead and/or zinc, including dross containing the metal in the form of an oxide or sulfide formed and removed in a subsequent processing step, a metal silicide, preferably a silicide of a metal selected from copper, zinc, nickel, iron, lead and tin, and a crust or other solid substance formed on the wall of a crucible or ladle that was used to move the molten metal and/or molten slag removed from the furnace. The inventors have found that the smelting stage is a very suitable place in the process for the recovery of by-products that may be rather poorly defined in terms of their content, such as some of the materials in the above list, or by-products that may contain various metals of interest, such as zinc oxide dust that can be collected by filtering the exhaust gases of furnaces performing a wide range of pyrometallurgical process operations, or furnace slags that contain recoverable metals in quantities that complicate or preclude their conventional recovery, and/or justify an additional pass through the entire metal recovery process.

In one embodiment of the method according to the invention, the raw material is introduced into the center of the liquid bath in the furnace. This has the advantage that the solid raw material, due to its buoyancy, which it experiences when immersed in the molten metal phase, can generally float on the surface of the molten metal phase without coming into contact with the refractory lining of the furnace. This reduces the wear of the refractory lining that the solid raw material can cause and, therefore, increases the service life of the refractory lining and the time between two maintenance operations for repairing the refractory lining.

In one embodiment of the method according to the present invention, the raw material comprises a coarse-grained portion, wherein the coarse-grained portion of the raw material preferably has an average particle size of at least 5 mm, preferably at least 10 mm, even more preferably at least 15 mm, and the average composition of the coarse-grained portion of the raw material fed to the entire batch of charge for melting in the furnace corresponds to at least one, and preferably all, of the following conditions after heating to 1150°C:

- contains at least 20% by weight of the total amount of metal, preferably at least 20% of the total amount of copper, nickel, tin, lead and zinc, preferably at least 30% by weight, more preferably at least 40% by weight, even more preferably at least 50% by weight, preferably at least 60% by weight, more preferably at least 70% by weight and optionally no more than 95% by weight of the total amount of metal, preferably of the total amount of copper, nickel, tin, lead and zinc,

- contains at least 10.0 wt.% and no more than 70.0 wt.% copper (Cu), preferably at least 15.0 wt.%, more preferably at least 17.0 wt.%, even more preferably at least 18.0 wt.%, even more preferably at least 19.0 wt.%, and optionally no more than 65.0 wt.%, preferably no more than 60.0 wt.%, more preferably no more than 55.0 wt.%, even more preferably no more than 50.0 wt.% and most preferably no more than 45.0 wt.% copper,

- contains at least 0.50 wt.% and no more than 2.00 wt.% nickel (Ni), preferably at least 0.60 wt.%, more preferably at least 0.70 wt.%, even more preferably at least 0.80 wt.%, even more preferably at least 0.90 wt.%, and optionally no more than 1.90 wt.%, preferably no more than 1.80 wt.%, more preferably no more than 1.70 wt.%, even more preferably no more than 1.60 wt.% and most preferably no more than 1.50 wt.% nickel,

- contains at least 1.00 wt.% and no more than 8.00 wt.% lead (Pb), preferably at least 1.10 wt.%, more preferably at least 1.25 wt.%, even more preferably at least 1.50 wt.%, even more preferably at least 1.60 wt.%, and optionally no more than 7.50 wt.%, preferably no more than 7.00 wt.%, more preferably no more than 6.50 wt.%, even more preferably no more than 6.00 wt.% and most preferably no more than 5.50 wt.% lead,

- contains at least 0.50 wt.% and at most 2.50 wt.% tin (Sn), preferably at least 0.60 wt.%, more preferably at least 0.70 wt.%, even more preferably at least 1.00 wt.%, even more preferably at least 1.20 wt.%, and optionally at most 2.40 wt.%, preferably at most 2.30 wt.%, more preferably at most 2.20 wt.%, even more preferably at most 2.00 wt.% and most preferably at most 1.90 wt.% tin,

- contains no more than 0.10 wt.% antimony (Sb), preferably no more than 0.08 wt.%, and more preferably no more than 0.06 wt.% antimony,

- contains at least 10.0 wt.% and no more than 35.00 wt.% iron (Fe), preferably at least 11.0 wt.%, more preferably at least 12.0 wt.%, even more preferably at least 13.0 wt.%, even more preferably at least 14.0 wt.%, and optionally no more than 34.5 wt.%, preferably no more than 34.0 wt.%, more preferably no more than 33.0 wt.%, even more preferably no more than 32.0 wt.% and most preferably no more than 31.0 wt.% iron, and

- contains at least 2.00 wt.% and no more than 15.00 wt.% zinc (Zn), preferably at least 2.50 wt.%, more preferably at least 3.00 wt.%, even more preferably at least 3.50 wt.%, even more preferably at least 4.00 wt.%, and optionally no more than 14.00 wt.%, preferably no more than 12.00 wt.%, more preferably no more than 11.00 wt.%, even more preferably no more than 10.00 wt.% and most preferably no more than 9.00 wt.% zinc.

The inventors have found that the coarse portion of the feedstock is very suitable as a feedstock for the process according to the present invention. This portion contains a sufficient amount of the metals of interest to make it of interest overall, but the levels of the valuable metals are not high enough to make it of interest for alternative processes for the recovery of some of these metals. The inventors have found that such a coarse portion is not rich enough in copper and/or tin and lead to make this portion a suitable feedstock for pyrometallurgical copper refining, as described, for example, in patent document WO 2019/115533 A1. Compliance with the upper limit of the sulfur content within one of the conditions in the above list further helps to avoid the formation of a separate copper matte phase, and thus clearly distinguishes the process of the present invention incorporating this feature from copper smelting processes in which the matte phase is formed as one of the products or intermediates.

In addition, the inventors have found that relatively low levels of sulphur are quite tolerable in the coarse portion of the feedstock. This has the advantage that a wider range of feedstocks can be used in the smelting stage, including feedstocks that would be unacceptable or less desirable in alternative methods of processing such feedstocks.

In one embodiment of the method according to the present invention, the content of iron and/or metals and compounds, at most as noble as iron or zinc, dissolved in the molten metal in the furnace is maintained at a level of at least 1.0 wt.%, preferably at least 1.5 wt.%, wherein the concentration of metals and compounds, at most as noble as iron or zinc, is converted into an equivalent concentration of iron, wherein the equivalent concentration of iron is the concentration of iron, which is capable of contributing the same amount of heat of reaction in the form of a metal or compound, not more noble than iron or zinc, by reacting with oxygen under the conditions of the furnace. The inventors have found that compliance with this condition is very convenient to control and maintain, and it easily ensures a sufficient excess of the elemental form of iron and metals or compounds, which under the conditions of the furnace are at most as noble as iron or zinc. Compliance with this condition also ensures that there is always a sufficient excess of iron and/or other metals or compounds, at most as noble as iron or zinc, in the furnace so that with a sufficient oxygen supply the temperature inside the furnace can be easily maintained. An additional advantage is that this condition ensures that the tuyere is protected by solid iron and/or iron oxide, as described elsewhere in this document.

In one embodiment of the method according to the present invention, the level of iron and/or metals and compounds that are at most as noble as iron or zinc under furnace conditions dissolved in the molten metal in the furnace is maintained at a level of at most 10.0 wt.%, preferably at most 9.0 wt.%, more preferably at most 8.0 wt.%, even more preferably at most 7.0 wt.%, even more preferably at most 6.0 wt.%, preferably at most 5.0 wt.%, more preferably at most 4.0 wt.%, even more preferably at most 3.5 wt.%, even more preferably at most 3.0 wt.%, preferably at most 2.5 wt.%, wherein the concentration of metals and compounds that are at most as noble as iron or zinc is converted into an equivalent concentration of iron, wherein the equivalent concentration of iron is the concentration of iron that is capable of contributing the same amount of heat of reaction in the form of a metal or compound that is at most as noble as iron or zinc by reacting with oxygen in furnace conditions. This reduces the risk of iron coming out of solution in colder areas of the furnace, such as near the furnace walls, where it can reduce the available furnace volume and impair mixing of the liquid bath inside the furnace.

In one embodiment of the method according to the present invention, elemental iron is introduced into the smelting stage at a rate that maintains an excess of iron in the smelting bath during the method, exceeding its solubility in the metal bath under the furnace conditions. The inventors have found that this is a very convenient means for providing a sufficient amount of iron in the furnace to ensure the desired excess.

In one embodiment of the method according to the present invention, the amount of excess iron present in the furnace is maintained by at least periodically taking samples of the molten metal phase in the furnace and analyzing them for the presence of iron. Preferably, the amount of excess iron is kept limited in order to keep the amount of solid iron parts moving in the liquid bath of the furnace limited in order to limit the possible damage and wear that these floating parts can cause to the refractory lining in the furnace.

In one embodiment of the method according to the present invention, the combustible source of carbon and/or hydrogen is selected from the group consisting of coke, charcoal, soot, hydrocarbon, natural gas, methane, ethane, propane, butane, hydrocarbon which is liquid under normal conditions, hydrocarbon-containing polymer, plastic, waste plastic, fat, oil, paint, varnish, rubber, preferably waste thereof, and combinations thereof. The inventors have found that a wide range of sources are suitable, and some of these sources are quite affordable on attractive delivery terms.

In one embodiment of the method according to the invention, the amount of combustible carbon and/or hydrogen source is maintained below a level at which slag foaming may impair the performance of the smelting stage, preferably significantly below this level so that the risk of slag foaming remains acceptably low. The inventors have found that the upper acceptable level depends on the source selected, but can be easily determined by trial and error. A further advantage of observing this precaution is that the temperature of the furnace exhaust gases remains acceptable, as does the carbon monoxide content of these exhaust gases. Alternatively, the amount of combustible carbon and/or hydrogen source is kept limited below a level at which the temperature of the exhaust gases remains acceptably low, or below a level at which the carbon monoxide content of these exhaust gases remains acceptably low.

In one embodiment of the method according to the present invention, at least a portion of the oxygen-containing gas is introduced into the upper phase of the slag, preferably as close as possible to the interface between the metallic phase and the slag phase. The inventors prefer to introduce at least a portion of the oxygen-containing gas at this target position, where oxygen is most easily consumed in the oxidation of elemental metals, such as iron, dissolved in the metallic phase, and from which the oxide formed as a result of the oxidation reaction can easily pass into the upper phase of the slag with a minimum diffusion distance that it must travel.

In one embodiment of the method according to the present invention, at least part of the oxygen-containing gas is introduced by means of at least one metal tuyere, the tip of which is immersed in the liquid phase of the slag. The inventors have found that this is a very convenient way of feeding the oxygen-containing gas to the desired position. The tuyere can be introduced through a special opening in the wall of the furnace, or can be introduced through the neck of the furnace, through which the raw material can also be introduced.

In one embodiment of the method according to the invention, using at least one metal tuyere, the gas introduced through the metal tuyere contains at least 30 vol.% oxygen, preferably at least 40 vol.%, more preferably at least 50 vol.%, even more preferably at least 75 vol.%, and most preferably this gas is high-purity oxygen. This has the advantage, compared to using air as the oxygen-containing gas, of reducing and, preferably, preventing the formation of additional exhaust gases. Thus, the exhaust gas treatment system can be reduced or can operate more efficiently. A further advantage is that the exhaust gases of the furnace contain less nitrogen oxides, and are therefore more environmentally acceptable.

In one embodiment of the method according to the present invention, using at least one metal tuyere, the gas flow through the metal tuyere provides sufficient cooling to prevent corrosion and/or melting of the tuyere under conditions of immersion in a liquid bath of molten slag. The inventors prefer to introduce the gas tuyere into the furnace from above the liquid bath, and to immerse the tuyere only in the liquid slag phase, but not in the bath of molten metal. The inventors have noticed that the tuyere, due to the sufficient cooling effect of the gas passing through it, as prescribed, can withstand a long exposure to the hot slag phase, but dissolves quite quickly in the underlying molten metal phase.

In one embodiment of the method according to the present invention, at least part of the oxygen-containing gas is introduced into the lower part of the furnace through at least one tuyeres, preferably through a plurality of tuyeres, more preferably uniformly distributed over the bottom of the furnace. This gives the advantage of strong mixing of the liquid bath in the furnace.

In one embodiment of the method according to the present invention using a tuyere, the gas introduced through at least one tuyere is an oxygen-containing gas containing no more than 50 vol.% oxygen, preferably no more than 40 vol.%, more preferably no more than 30 vol.%, even more preferably no more than 25 vol.%, and most preferably the gas introduced by the tuyere is air. This gas must overcome the hydrostatic pressure caused by the height of the liquid contents of the furnace. Therefore, the gas must be compressed so that it can be introduced through the tuyere. If the gas contains air, this air is compressed before being introduced.

In one embodiment of the method according to the present invention, using a tuyere, the gas introduced through at least one tuyere is cooler than the molten liquid metal phase surrounding the tuyere. The molten liquid metal phase around the tuyere is thereby cooled, and the solubility of iron in it is thereby reduced, and when the molten liquid metal phase is saturated with iron at a higher temperature, this leads to iron and/or iron-containing compounds, such as iron oxides, being deposited and deposited around the tuyeres, usually in the form of a hollow mushroom, which provides reliable protection of the tuyeres from corrosion caused by the high oxidation temperature of the iron near the tuyeres.

In one embodiment of the method according to the present invention, the smelting slag produced by the method contains at least 20 wt.% iron (Fe), preferably at least 22.5 wt.%, more preferably at least 25.0 wt.%, even more preferably at least 27.50 wt.%, and most preferably at least 30.00 wt.% iron. In this context, the iron content is the sum of the iron content in all its valence states, i.e. the sum of all the iron present as elemental iron and the iron present in a chemically bound form, usually in the form of an oxide. This gives the advantage of higher slag fluidity, i.e. a lower viscosity at the same temperature.

In one embodiment of the method according to the present invention, the composition of the produced smelting slag meets at least one, and preferably all, of the following conditions:

- contains no more than 1.00 wt.% copper (Cu), preferably no more than 0.90 wt.%, more preferably no more than 0.80 wt.%, even more preferably no more than 0.70 wt.%, and even more preferably no more than 0.60 wt.% copper,

- contains no more than 0.20 wt.% nickel (Ni), preferably no more than 0.17 wt.%, more preferably no more than 0.15 wt.%, even more preferably no more than 0.12 wt.%, and even more preferably no more than 0.10 wt.% nickel,

- contains no more than 2.00 wt.% lead (Pb), preferably no more than 1.50 wt.%, more preferably no more than 1.00 wt.%, even more preferably no more than 0.95 wt.%, and even more preferably no more than 0.90 wt.% lead,

- contains no more than 1.00 wt.% tin (Sn), preferably no more than 0.80 wt.%, more preferably no more than 0.60 wt.%, even more preferably no more than 0.40 wt.%, and even more preferably no more than 0.25 wt.% tin,

- contains no more than 22.50 wt.% zinc (Zn), preferably no more than 20.00 wt.%, more preferably no more than 17.50 wt.%, even more preferably no more than 15.00 wt.%, and even more preferably no more than 12.50 wt.% zinc.

Compliance with the upper limits specified for copper, nickel, tin and lead provides the advantage of a low yield of valuable metals from the process. Since the present invention is directed to a copper recovery process, limiting copper losses in the smelter slag provides the advantage of a high copper recovery from the available raw materials.

The inventors have found that many copper-containing raw materials, particularly the secondary materials in this group, also contain significant amounts of primarily tin, but possibly also lead, nickel and zinc. The inventors have found that most of these metals, except copper, can be recovered from the same raw materials by pyrometallurgical process steps, provided that these metals do not escape with the smelter slag. Adherence to upper limits for the other metals, primarily tin and nickel, but also lead and to some extent also zinc, gives the advantage of high recovery of these metals from the available raw materials.

The presence of zinc in the smelting slag may be increased if an additional slag evaporation step is provided, in which the smelting slag is evaporated to reduce its zinc content and optionally also its lead content. The inventors prefer to add such an additional evaporation step to remove more zinc from the smelting slag, preferably also additional traces of lead, preferably as described in patent document WO 2016/156394 A1.

The inventors have found that the above-mentioned features of low losses of valuable metals in the smelting slag can be controlled and can be obtained by suitable operation of the smelting stage in terms of furnace temperature, furnace stirring and oxygen addition, reducing agents and their selection, as well as flux (or slag former) materials and their selection.

In one embodiment of the method according to the present invention, the composition of the concentrated intermediate copper product, which is the main product of the smelting stage, meets at least one, and preferably all, of the following conditions:

- contains at least 50.0 wt.% copper (Cu), preferably at least 55.0 wt.%, more preferably at least 60.0 wt.%, even more preferably at least 65.0 wt.%, even more preferably at least 70.0 wt.%, preferably at least 72.5 wt.%, more preferably at least 75.0 wt.%, even more preferably at least 77.0 wt.%, even more preferably at least 78.0 wt.% or even 79.0 wt.% copper (Cu), and optionally at most 97.0 wt.%, preferably at most 95.0 wt.%, more preferably at most 90.0 wt.%, even more preferably at most 85 wt.%, even more preferably at most 82.0 wt.%, preferably at most 80 wt.%, more preferably at most 79.0 wt.%, even more preferably at most 78.0 wt.%, and even more preferably not more than 77.0 wt.% copper (Cu),

- contains at least 0.01 wt.% nickel (Ni), preferably at least 0.05 wt.%, more preferably at least 0.10 wt.%, even more preferably at least 0.50 wt.%, even more preferably at least 1.00 wt.%, preferably at least 1.10 wt.%, more preferably at least 1.25 wt.%, even more preferably at least 1.40 wt.%, even more preferably at least 1.50 wt.% or even 1.70 wt.% nickel (Ni), and optionally at most 15.00 wt.%, preferably at most 12.50 wt.%, more preferably at most 10.00 wt.%, even more preferably at most 7.50 wt.%, even more preferably at most 5.00 wt.%, preferably at most 4.00 wt.%, more preferably at most 3.00 wt.%, even more preferably not more than 2.50 wt.%, and even more preferably not more than 2.40 wt.% nickel (Ni),

- contains at least 0.10 wt.% lead (Pb), preferably at least 0.50 wt.%, more preferably at least 1.00 wt.%, even more preferably at least 2.00 wt.%, even more preferably at least 3.00 wt.%, preferably at least 3.50 wt.%, more preferably at least 4.00 wt.%, even more preferably at least 4.50 wt.%, even more preferably at least 5.00 wt.% or even 5.50 wt.% lead, and optionally at most 15.00 wt.%, preferably at most 14.50 wt.%, more preferably at most 14.00 wt.%, even more preferably at most 13.50 wt.%, even more preferably at most 13.00 wt.%, preferably at most 12.50 wt.%, more preferably at most 12.00 % by weight, even more preferably not more than 11.50 % by weight, even more preferably not more than 11.00 % by weight, preferably not more than 10.50 % by weight, more preferably not more than 10.00 % by weight, even more preferably not more than 9.50 % by weight, and even more preferably not more than 9.00 % by weight of lead (Pb),

- contains at least 1.00 wt.% tin (Sn), preferably at least 1.25 wt.%, more preferably at least 1.50 wt.%, even more preferably at least 1.75 wt.%, even more preferably at least 2.00 wt.%, preferably at least 2.25 wt.%, more preferably at least 2.50 wt.%, even more preferably at least 2.75 wt.%, even more preferably at least 3.00 wt.% or even 3.25 wt.% tin (Sn), and optionally at most 12.00 wt.%, preferably at most 10.00 wt.%, more preferably at most 8.00 wt.%, even more preferably at most 7.00 wt.%, even more preferably at most 6.00 wt.%, preferably at most 5.50 wt.%, more preferably at most 5.00 wt.%, even more preferably not more than 4.50 wt.%, and even more preferably not more than 4.00 wt.% tin (Sn),

- contains at least 0.05 wt.% iron (Fe), preferably at least 0.10 wt.%, more preferably at least 0.30 wt.%, even more preferably at least 0.50 wt.%, even more preferably at least 0.60 wt.%, preferably at least 0.70 wt.%, more preferably at least 0.80 wt.%, even more preferably at least 0.90 wt.%, even more preferably at least 1.00 wt.% or even 1.10 wt.% iron (Fe), and optionally at most 5.00 wt.%, preferably at most 4.00 wt.%, more preferably at most 3.00 wt.%, even more preferably at most 2.50 wt.%, even more preferably at most 2.00 wt.%, preferably at most 1.75 wt.%, more preferably at most 1.50 wt.%, even more preferably at most 1.25 wt.%, and even more preferably not more than 1.00 wt.% iron (Fe),

- contains at least 0.10 wt.% zinc (Zn), preferably at least 0.50 wt.%, more preferably at least 1.00 wt.%, even more preferably at least 2.00 wt.%, even more preferably at least 2.50 wt.%, preferably at least 3.00 wt.%, more preferably at least 3.50 wt.%, even more preferably at least 4.00 wt.% zinc (Zn), and optionally at most 10.00 wt.%, preferably at most 9.50 wt.%, more preferably at most 9.00 wt.%, even more preferably at most 8.50 wt.%, even more preferably at most 8.00 wt.%, preferably at most 7.50 wt.%, more preferably at most 7.00 wt.%, even more preferably at most 6.50 wt.%, even more preferably at most 6.00 wt.%, preferably not more than 5.50 wt.%, and more preferably not more than 5.00 wt.% zinc (Zn),

- contains at most 5 wt.% sulfur (S), preferably at most 4.5 wt.%, more preferably at most 4.0 wt.%, even more preferably at most 3.5 wt.%, even more preferably at most 3.0 wt.%, preferably at most 2.5 wt.%, more preferably at most 2.0 wt.%, preferably at most 1.5 wt.%, more preferably at most 1.0 wt.%, even more preferably at most 0.5 wt.%, even more preferably at most 0.1 wt.% sulfur, and optionally at least 5 ppm, preferably at least 50 ppm, more preferably at least 100 ppm, even more preferably at least 500 ppm, even more preferably at least 1000 ppm per million, preferably at least 0.5 wt.%, and more preferably at least 1.0 wt.% sulfur.

The inventors have found that the above-described features of the present invention can also be controlled and obtained by appropriately performing the smelting step as described above, including the selection of feed materials. Maintaining an upper limit of sulfur content within one of the conditions in the above list further avoids the formation of a separate copper matte phase, and thus clearly distinguishes the process of the present invention, incorporating this feature, from copper smelting processes in which the matte phase is formed as one of the products or intermediate products. The inventors have also found that the above-described metallic phase is very suitable for the recovery of valuable metals listed in the stages of the pyrometallurgical process described below in this document.

The inventors also found that a large amount of copper in the metallic phase can act as an extractive agent for the extraction of other valuable metals such as nickel, tin and lead from the slag phase and, therefore, in itself also contributes to a high recovery of these metals other than copper.

The inventors have also found that maintaining the specified lead content provides advantages in the recovery of even more economically viable metallic tin, since tin and lead can be recovered as a solder-type by-product stream, which, due to the lead content, can be suitably adjusted, and then distilled to recover a high-purity tin product together with lead-containing by-products of varying quality, which are also valuable.

In one embodiment of the method according to the present invention, the liquid bath in the furnace has a temperature in the range of 1100-1300°C, preferably at least 1120°C, more preferably at least 1140°C or even 1150°C, and optionally not more than 1250°C, preferably not more than 1200°C, more preferably not more than 1180°C. The inventors have found that this temperature range can ensure sufficient slag fluidity and adequate zinc evaporation already in the smelting stage, while maintaining a low evaporation rate of tin and/or lead in the smelting stage, which therefore contributes to a high recovery of tin and/or lead and/or zinc, as well as a high productivity of the smelting stage. The inventors prefer to allow the contents of the furnace to cool to not more than 1140°C before the metal is removed from the furnace. The inventors have found that this precaution helps to increase the service life of the containers in which the molten metal is received and transferred to the subsequent stage of the process.

In one embodiment of the method according to the present invention, the off-gases from the furnace are collected and treated by cooling and/or filtering. The inventors have found that the off-gases from the smelting stage contain valuable metals that are useful to recover, and that said treatment also reduces environmental problems associated with the emission of off-gases from the smelting stage into the atmosphere.

In one embodiment of the method according to the present invention, secondary exhaust gases from the entire furnace are also collected and treated by filtration, optionally in combination with cooling. The inventors have found that this feature further reduces environmental problems that may be associated with the operation of the furnace according to the present invention.

In one embodiment of the method according to the present invention, the melting is carried out in a melting furnace.

The smelting furnace offers the advantage of simplicity in operation and equipment, and is therefore economically advantageous. The smelting furnace offers the additional advantage of being tolerant of the quality of the raw material. The smelting furnace is able to accept raw materials that are extremely diluted and/or contaminated with various components, such as various organic substances. In the smelting furnace, the metals are melted and the organics and other combustible materials are burned. Since this mixed and/or contaminated raw material is unlikely to have any other end use, it can be supplied on economically very attractive terms. Therefore, the ability to process such raw materials and to recover the valuable metals contained therein is of interest to the operator of the method according to the present invention.

The melting furnace is a relatively simple and cheap device consisting of a large cylindrical furnace that only needs to be able to tilt around its longitudinal axis by a fraction of a full circle. This discovery provides the advantage of low capital investment and/or operating costs for the melting stage.

In one embodiment of the method according to the invention, the furnace walls are cooled at least partially on their surface. This has the advantage of reduced wear on the furnace wall, and in particular on the moving parts provided as part of the means for moving the furnace for stirring the liquid bath, and for improving and/or regulating the stirring of the bath.

In one embodiment of the present invention, the process further comprises a step of fuming the slag phase formed in the smelting step to produce a evaporated slag, and preferably the fuming is carried out in a fuming furnace. The fuming step produces an evaporated slag, at least together with dust, containing most of the metals evaporated from the smelting slag, usually in their oxidized form. The inventors have found that it is advantageous to provide this additional process step, since it expands the acceptance criteria of the raw material for the smelting step by including raw materials that contain more zinc, and optionally also lead. Such raw materials are often difficult to process in alternative processes, where they can create an unnecessary technological and/or economic burden, and therefore they can be available in greater quantity and on more attractive economic terms. The zinc evaporation step can be performed as described in Michael Borell, “Slag - a resource in the sustainable society”, proceedings of the International Conference on Mining and the Environmental Metals and Energy Recovery “Securing the Future”, Sweden, 2005, pp. 130-138. The inventors, however, prefer to perform such an additional fuming step as disclosed in patent document WO 2016/156394 A1.

In one embodiment of the method according to the present invention, the slag is granulated when removed from the smelting stage or from the fuming stage. Preferably, the slag from the smelting stage and/or the slag from the evaporation stage is removed from the respective furnaces in the form of a liquid. This is an advantage in that the furnace can be emptied for further production and/or processing of the slag while the resulting slag cools. The slag can be cooled and/or solidified by contacting the slag with a cooling agent, such as air, possibly air from the environment.

In one embodiment of the method according to the present invention, slag cooling is performed by contacting the liquid slag with water. The inventors have found that water cooling is very effective and can be applied in a variety of ways, resulting in relatively well-controlled cooling rates.

In one embodiment of the method according to the present invention, the process further comprises the step of using the produced slag for an end use selected from providing a wear layer and/or coating for roofing tiles or slates, a component for sandblasting, a foamed component of tiles, a black dye, preferably in building products, more preferably in black tiles, in the form of black solid pieces, preferably for decorative purposes, and as a high-density ballast, preferably for underwater applications, more preferably for hydraulic engineering, as well as combinations thereof.

In one embodiment of the method according to the present invention for producing an object for the construction industry, the process further comprises a step of adding the produced slag as an aggregate and/or a binder during the production of the object for the construction industry, preferably as a binder for aggregates, preferably as an active binder, more preferably as a binder having pozzolanic activity, even more preferably as a replacement for Portland cement, even more preferably as a partial replacement for Portland cement.

In one embodiment of the method according to the present invention, the produced slag is added as a binder to the inorganic polymer composition, preferably in combination with a base, more preferably as a main binder, and even more preferably as the sole binder in the inorganic polymer composition.

In one embodiment of the method according to the present invention, in which the produced slag is added during the production of an object for the construction industry, the process further comprises the step of foaming the inorganic polymer composition.

In one embodiment of the method according to the present invention for producing an object for the construction industry, the object for the construction industry is a building element, preferably a building element selected from a tile, a paving stone, a block, a concrete block and combinations thereof.

In one embodiment of the method according to the present invention for producing an object for the construction industry, the object for the construction industry has a foamed structure.

In one embodiment of the method according to the present invention for producing an object for the construction industry, the process further comprises the step of using the object to improve thermal and/or sound insulation, to shield X-ray radiation, as well as for combinations thereof.

In one embodiment of the method according to the present invention, the process further comprises a step of purifying the concentrated intermediate copper product to obtain a purified copper product together with at least one copper refining slag. The inventors have found that this purification step can be suitably performed as described in patent document WO 2019/115543 A1.

In one embodiment of the method according to the present invention, in which the concentrated intermediate copper product further comprises tin and lead, the process further comprises extracting a raw solder metal composition from the concentrated intermediate copper product. This extraction of the raw solder can suitably be carried out as described in patent document WO 2019/115524 A1.

In one embodiment of the method according to the present invention comprising recovering a metallic composition of green solder, the process further comprises a step of recovering from the metallic composition of green solder at least one of a purified soft lead product, a purified hard lead product and a purified tin product. The inventors have found that the metallic composition of green solder is a very suitable raw material for recovering at least one of the listed products, preferably at least the purified tin product, more preferably also one of the purified lead products, and even more preferably both purified lead products.

In one embodiment of the method according to the present invention, comprising recovering a raw solder metal composition, the process further comprises a step of pre-purifying the raw solder metal composition in order to produce a pre-purified solder metal composition.

The metallic composition of the green solder, which may be obtained as a by-product of refining the concentrated intermediate copper product obtained in the process according to the present invention, may further be pre-cleaned or treated to remove therefrom a greater amount of impurities, in particular copper. This may be accomplished by contacting the metallic composition of the green solder in the form of a molten liquid with elemental silicon and/or aluminum, elements which, under the operating conditions, combine with Cu, Ni and/or Fe and form a separate phase of the silicide and/or aluminide alloy. The inventors prefer to use silicon- and/or aluminum-containing waste. Preferably, the added material further comprises Sn and/or Pb, because these metals are easily converted to the corresponding main products when introduced at this stage of the process. Due to the typical presence of Sb and As in the metallic composition of the green solder, the inventors prefer to use silicon and avoid aluminum, although it is usually more readily available and more reactive. This avoids the formation of _H2S , a toxic gas, and more exothermic reactions in the processing vessel, and also prevents the resulting alloy phase by-product from forming stibine and/or arsine, highly toxic gases, upon contact with water. The inventors have found that the silicon feed for this processing step can contain a limited amount of iron (Fe), greater than 1 wt.% and as high as 5 wt.% or even as high as 10 wt.% Fe. The process can thus operate using Si products that are unacceptable to other silicon consumers, such as manufacturing waste, and which can thus be more readily available. The inventors have found that the cost of processing this additional Fe, which also binds to Si, is usually easily offset by the favorable supply conditions of the silicon source.

This pre-cleaning can suitably be carried out as described in patent document WO 2019/115524 A1, to obtain a by-product of the so-called "cupro phase", which can preferably be returned to the smelting stage of the process according to the present invention after the described "washing".

In one embodiment of the method according to the present invention, comprising recovering a metallic composition of the green solder, the process further comprises a step of adjusting the metallic composition of the green solder or the pre-cleaned solder metal composition to obtain an adjusted solder metal composition. This adjustment step can further prepare the solder so that it becomes suitable for vacuum distillation, a technically very demanding step of the method, which is sensitive to the excessive presence of certain metallic impurities. Such adjustment and distillation can suitably be carried out as described in patent document WO 2018/060202 A1.

In one embodiment of the method according to the present invention, producing an adjusted solder metal composition, the process further comprises a step of first distillation of the adjusted solder composition, in which lead is removed from the solder by evaporation, and a first distillation top product and a first distillation bottom product are obtained. Such a first distillation can suitably be carried out as described in patent document WO 2018/060202 A1.

In one embodiment of the method according to the present invention producing a first distillation overhead product, the process further comprises a step of removing at least one impurity selected from the metals arsenic, antimony and tin from the first distillation overhead product to obtain a purified soft lead product. Preferably, the purified soft lead product is produced as described in patent document WO 2020/157165 A1.

In one embodiment of the method according to the present invention, comprising a first distillation step, the bottom product of the first distillation step comprises lead and silver, and the process further comprises the step of separating the bottom product of the first distillation step by fractional crystallization into a first silver-rich liquid product and a first tin-rich crystalline product. The inventors prefer to carry out this separation as described in patent document WO 2020/157167 A2.

In one embodiment of the method according to the present invention, comprising a fractional crystallization step, the process further comprises a step of separating the first silver-rich liquid product into a product rich in lead and tin and a product rich in silver, preferably by electrolysis, in which the anode sludge is a product rich in silver. The inventors prefer to carry out this separation as described in patent document WO 2020/157167 A2.

In one embodiment of the method according to the present invention, producing a first tin-rich product, and this first tin-rich product further comprises lead and antimony, the process further comprises a second distillation step of the first tin-rich product, in which the lead and antimony are mainly evaporated, and a second distillation overhead product and a second distillation bottom product are obtained.

In one embodiment of the method according to the present invention, comprising a second distillation step, the process further comprises a third distillation step of the top product of the second distillation, in which the lead is evaporated and the top product of the third distillation and the bottom product of the third distillation are obtained, and preferably the bottom product of the third distillation is at least partially returned to the input of the second distillation step and/or to the input of the fractional crystallization step.

In one embodiment of the method according to the present invention comprising a third distillation step, the process further comprises a step of removing at least one impurity selected from the metals arsenic and tin from the overhead product of the third distillation to obtain a purified solid lead product. The inventors prefer to perform this step as described in patent document WO 2020/157168 A1.

In one embodiment of the method according to the present invention comprising a second distillation step, the process further comprises a step of purifying the bottom product of the second distillation to obtain a purified tin product. The inventors prefer to perform this step as described in patent document WO 2020/157168 A1.

In one embodiment of the method according to the present invention, comprising purifying a concentrated intermediate copper product to obtain a purified copper product, the process further comprises a step of casting the purified copper product to obtain purified copper anodes. The inventors have found that the purified copper product in the form of copper anodes is very suitable for a further step of the electrolytic process for producing high-purity copper cathodes together with anode slimes that can be further processed to recover the valuable metals contained therein. The inventors prefer to perform this step of the electrolytic purification process as described in patent document WO 2019/219821 A1.

In one embodiment of the method according to the present invention, at least part of the method is monitored and/or controlled electronically. The inventors have found that controlling the stages of the method according to the present invention electronically, preferably by a computer program, gives the advantage of a much better processing, with much more predictable results that are closer to the objectives of the method. For example, based on temperature measurements, and optionally also pressure and/or level measurements, and/or in combination with the results of chemical analyses of samples taken from the process streams and/or with the results of analysis obtained in real time, the control program can control equipment related to switching on or off electrical power, supply of heat or cooling agent, flow and/or pressure regulation. The inventors have found that such monitoring or control is particularly advantageous with stages that are operated in a continuous mode, but it can also be advantageous with stages that are operated in a batch or semi-continuous mode. In addition to this and preferably, the results of the monitoring obtained during or after the execution of the stages in the process according to the present invention are also used for monitoring and/or controlling other stages that are part of the method according to the present invention and/or processes that are applied before or after the method according to the present invention, as part of an overall method of which the process according to the present invention is only a part. Preferably, the entire overall process is monitored electronically, more preferably by at least one computer program. Preferably, the overall method is controlled electronically to the greatest possible extent.

The inventors prefer that the computer control also provides for the transmission of data and instructions from one computer or computer program to at least one other computer or computer program, or to a module of the same computer program, for monitoring and/or controlling in other ways, including, but not limited to, the processes described herein.

The present invention is further illustrated in Fig. 1, which shows a flow chart of the entire process as a preferred embodiment, comprising the steps of the process according to claim 1 for recovering the concentrated intermediate copper product.

In Fig. 1, the following reference numerals refer to the following process steps or flows:

100. Melting stage or melting furnace

200. Cleaning copper

300. Casting copper anodes

400. Slag evaporation stage

500. Extraction of lead/tin

1. Coarse-grained part of raw material

2. Finely ground raw material or dust

3. Blister copper as a concentrated intermediate copper product

4. By-product of smelter dust

5. Melting slag

6. Refining slag

7. A by-product of copper refining in the form of raw solder

8. Refined copper

9. Product in the form of copper anodes

10. Product in the form of soft lead

11. Product in the form of solid lead

12. Product in the form of purified tin

13. Evaporated slag

14. By-product of the fuming furnace in the form of dust

Fig. 1 shows that a portion 1 of coarse raw material and a portion 2 of fine raw material are fed into a smelting furnace 100, where an oxygen-containing gas (not shown) is introduced to regulate the reactions in the furnace and thus also the temperature in the furnace. The exhaust gases from the furnace are cooled and filtered, as a result of which smelting dust 4 is collected. The smelting slag 5 is removed from the furnace and fed to a fuming stage 400 to extract the fuming dust 14 and to obtain a final slag or so-called "clean slag" 13 as a second slag.

Black copper 3 as a concentrated intermediate copper product is fed to copper purification 200, which produces purified copper product 8, by-product 7 raw solder and refining slag 6. Refining slag 6 can be sent to fuming furnace 400 to increase the amount of final slag 13 and fuming dust 14. Purified copper 8 is fed to casting stage 300 for producing copper anodes 9. Raw solder 7 is sent to lead/tin recovery stage 500, which produces soft lead product 10, purified tin product 12 and optionally hard lead product 11.

The inventors have found that the beneficial technical effects of the present invention, in particular the more stable and reliable operation of the smelting stage 100, are easily transferred downstream to the entire production of derivatives 9, 10, 11, 12, 13 and 14 shown in Fig. 1. The stages of obtaining derivatives, thanks to the present invention, have a more stable and reliable flow of raw materials coming from the smelting stage, and therefore can produce final products having a more stable and reliable quality. An additional advantage is that the present invention reduces the load on control and the attention of the operator to control the subsequent stages of the method, as well as the entire method as a whole.

EXAMPLE

The rotary drum furnace, which had an internal diameter of 3 m, maintained a level of approximately 1.00 m3 of liquid black copper as a concentrated intermediate copper product from the previous charge, which was approximately 113 metric tons.

The kiln was operated in a semi-continuous mode for a period of approximately 16 months using a repetitive sequence of the following operating modes, with each cycle relating to different batches of charge that were composed and assembled by selecting batches from a large stock of available raw materials:

Mode 1: Coarse-grained solid raw material is gradually fed into the furnace from a suitable batch of feed. This mode is switched on as needed until a continuous layer of slag is obtained floating on top of the liquid metal phase,

Mode 1+2: In the presence of a continuous slag layer in the furnace from another suitable batch of charge composed for this purpose, finely ground raw material, also called "dust", is transported pneumatically and blown into the liquid slag phase above the metallic phase of the liquid bath, usually also gradually, while the gradual feeding of coarse-grained solid material preferably also continues,

Mode 3: As a rule, but only if necessary, after the loading of raw materials is completed and/or the furnace is considered full, a period begins during which the conditions in the furnace are maintained for chemical reactions to occur until the desired slag and metal composition is obtained.

Mode 4: The slag is poured out of the furnace by tilting the rotating drum until the upper slag phase at least partially overflows through the furnace feed opening. The slag is preferably transferred in liquid form in a suitable container to a fuming furnace for further recovery of zinc and optionally also lead by evaporation and possibly also copper as part of the by-product of the metallic phase of the fuming stage. The slag obtained in this fuming stage or, if a fuming stage was not performed, the slag from the smelting stage, is cooled, solidified and granulated by contacting the hot liquid slag with a high flow of water.

Mode 5: The metal is partially removed from the furnace, if convenient, until a minimum level of about 1.00 m of liquid metal plus possibly some liquid slag remains in the furnace again. If the slag has been completely removed, removal is carried out by pouring the metal phase through the feed opening, and if only part of the slag phase has been removed, the metal is tapped through a suitably located taphole in the furnace wall.

After the slag and/or metal have been removed in Mode 4 and/or Mode 5, i.e. when there is space in the furnace again, the feed of raw materials is resumed depending on the presence of slag in the furnace, as in Mode 1 or Mode 1+2 described above, and if the previous batch of loading has ended, the next batch of loading begins.

If subsequent processing of the concentrated intermediate copper product requires more feedstock, a portion of the liquid metal phase is periodically removed from the furnace, either before or after removing any upper slag phase immediately before or immediately after metal removal.

Additional flux, also sometimes called "slag former" in translation of the term commonly used in German, usually sand, is added to the furnace as needed to ensure sufficient slag flow. Before granulating the slag from the furnace or from the subsequent fuming furnace, additional silicon dioxide is added if necessary to ensure the correct Fe/Si ratio and to eliminate the risk of hydrogen formation during pouring and granulation of the slag and the associated risk of explosion. During the period required to process a batch of charge, an average of 11.5 tons of sand per batch of charge was added to the furnace.

In modes 1, 1+2 and 3, when necessary to maintain the furnace temperature, pure gaseous oxygen was blown through a tuyere that was introduced through the inlet. In these operating modes, the rotary kiln was rocked whenever possible to mix its liquid contents.

In all operating modes, compressed air at a pressure of 10 bar was supplied to 4 tuyeres located at suitable positions in the furnace wall below the liquid level and was injected into the furnace, primarily for the purpose of stirring the bath, but also to introduce additional oxygen into the bath to participate in the intended chemical reactions.

During the period under review, an average of about 92.4 tons of coarse solids, including return, and about 23.2 tons of fines were introduced into the furnace per charge. The coarse solids and fines had the average composition shown in Table I. A sufficient amount of additional solid scrap iron was included in the pre-mixed coarse solids and fed into the furnace as part of it to maintain the presence of solid iron floating on the metallic phase. This amount of additional scrap iron introduced is thus included in the composition of the coarse solids shown in Table I.

Table I

Element (wt%) Coarse fraction Dust Cu 40,45 44.82 Pb 5.02 6.74 Sn 1.67 4.93 Ni 0.99 0.76 Fe 27.10 3.43 Si 1.99 1.70 Al 0.25 0,00 Zn 9.30 16.07 Bi 0.02 0.01 As 0.01 0.07 Sb 0.01 0.16 CaO 0.51 4.64 _SiO2 5.49 5.58 _{Al2O3 ​} 1.33 1.75 Total (wt.%) 94.14 90.66

During the entire period of operation considered, an average of about 6.6 tons of oxygen were introduced per charge in the form of compressed air at ambient temperature at the bottom of the furnace through four tuyeres and in the form of gaseous oxygen through a tuyeres near the interface between the metallic phase and the upper slag phase. The furnace temperature could be maintained very conveniently and accurately within a narrow range of 1150-1180°C. Most importantly, the control of the oxygen supply made it possible to avoid temperature jumps above this range, so that evaporation of tin and/or lead could be minimized. Adjustment of the feed rates of the raw material, including the coarse solid material, as well as the introduction of dust, made it possible to avoid a drop in temperature below the required level, and the introduction of oxygen made it possible to quickly recover from a temporary drop in temperature if such a drop occurred. For removal of metal from the furnace, the temperature was allowed to drop to about 1140°C in order to reduce the risk of damage to the containers into which the molten metal was poured.

The exhaust gases from the furnace were cooled and filtered to extract solids from the cooled gases in the form of smelter dust.

Over the entire period of operation under consideration, the average number of products and compositions obtained and removed from the smelting furnace per loading was shown in Table II.

Table II

Element (wt%) Smelting slag Metal Dust Cu 0.60 77.41 6.18 Pb 0.30 9.40 10.99 Sn 0.11 4.25 1.03 Ni 0.04 1.74 0.03 Fe 35.85 1,12 0.50 Si 0,00 0,00 0,00 Al 0,00 0,00 0,00 Zn 11.48 4.98 48.90 Bi 0,00 0.04 0,00 As 0,00 0.04 0.03 Sb 0,00 0.07 0.04 CaO 2.20 0,00 0,00 _SiO2 31.40 0.02 0.60 _{Al2O3 ​} 3.58 0,00 0.01 Total 85.56 99.07 68.31 Metric tons 70.1 61.1 2.5

The sulphur content of the metallic phase obtained as a concentrated intermediate copper product in the process described in the example was determined and was found to be significantly below 2 wt% for each batch of feed, but rather in the range of no more than 0.25 wt%. The sulphur content of the slag was found to be no more than 0.33 wt% for each batch of feed, and that of the dust no more than 0.21 wt%.

Having now fully described this invention, it will be apparent to those skilled in the art that the present invention may be carried out in a wide range of the stated parameters without departing from the scope of the invention as defined by the claims.

Claims (78)

1. A method for extracting copper from secondary raw materials, comprising the step of, in at least one batch of loading, smelting (100) a feedstock (1,2) containing raw materials in a furnace to extract from it a concentrated copper intermediate product (3), in which a feedstock of raw materials is gradually introduced into the furnace, the feedstock containing copper and, if present, at least one metal which, under the operating conditions of the furnace, has a lower affinity for oxygen than tin, at least partly in the form of an oxide, wherein the feedstock further comprises iron and, if present, at least one metal or compound that, under furnace conditions, has at least the same affinity for oxygen as iron or zinc, wherein the iron and the metal having at least the same affinity for oxygen as iron or zinc are at least partially present in elemental form, in which heat is generated inside the furnace as a result of oxidation-reduction reactions converting elemental iron and metals or compounds having at least the same affinity for oxygen as iron or zinc into oxides, and converting oxides of copper and metals having a lower affinity for oxygen than tin into the elemental metal, in which the elemental metals are at least partially collected in the molten liquid metal phase and the oxides are at least partially collected in the floating liquid slag phase, in which the separation of the liquid phases is ensured, and at the end of the smelting stage at least one of the liquid phases is at least partially removed from the furnace in the form of smelting slag (5) and/or in the form of a concentrated copper intermediate product (3), characterized in that during the smelting stage, the furnace is maintained in excess of elemental iron and, if present, metals or compounds which, under furnace conditions, have at least the same affinity for oxygen as iron or zinc, relative to the amount required to complete the oxidation-reduction reactions, and additional heat supply to the furnace is provided during the melting stage by feeding oxygen-containing gas to oxidize excess iron and metals or compounds having at least the same affinity for oxygen as iron or zinc present in the furnace and, if necessary, to burn the additionally introduced combustible source of carbon and/or hydrogen into the furnace. 2. The method according to claim 1, wherein the raw material additionally contains at least one second metal selected from the group consisting of nickel, tin and lead. 3. The method according to claim 2, wherein the concentrated intermediate copper product (3) additionally contains at least one second metal. 4. The method according to any one of paragraphs 1-3, wherein the raw material (1, 2) contains scrap iron, silicon, zinc and/or aluminum. 5. The method according to any one of claims 1 to 4, further comprising the step of removing the smelting slag (5) from the furnace, at least partially. 6. The method according to any one of claims 1 to 5, further comprising the step of removing from the furnace at least a portion of the concentrated intermediate copper product (3). 7. The method according to any one of claims 1 to 6, wherein the iron and compounds having at least the same affinity for oxygen as iron or zinc, which are introduced together with the raw material (1, 2), comprise solid iron, solid silicon, solid zinc and/or solid aluminium, preferably copper/iron-containing scrap, silicon-containing scrap, zinc-containing scrap, and/or aluminium-containing scrap. 8. The method according to any one of claims 1 to 7, wherein the raw material (1, 2) is at least partly solid, and wherein the solid raw material is gradually fed into the furnace, preferably continuously. 9. The method according to any one of claims 1 to 8, wherein the feedstock feed rate is maintained below the rate at which heat generation becomes insufficient to melt the solid feedstock and/or bring the feedstock to the desired furnace temperature. 10. The method according to any one of claims 1 to 9, wherein at least part of the raw material (1, 2) has the form of a finely ground part (2) which has an average particle size of no more than 10 mm. 11. The method according to claim 10, in which the finely ground material of part (2) of the raw material is transported pneumatically and introduced into the furnace. 12. The method according to claim 11, in which the material of the finely ground portion (2) of the raw material is introduced into the liquid slag phase above the metal phase of the liquid bath. 13. The method according to any one of claims 10 to 12, wherein the average composition of the finely ground portion (2) of the raw material fed to the entire batch of feed for melting in the furnace meets at least one, and preferably all, of the following conditions after heating to 1150°C: - contains at least 5% by weight of the total amount of metals, preferably at least 5% of the total amount of copper, nickel, tin, lead and zinc, - contains at most 70.0 wt.% copper (Cu), - contains at most 2.00 wt.% nickel (Ni), - contains at least 0.50 wt.% and at most 10.00 wt.% lead (Pb), - contains at most 15.00 wt.% tin (Sn), - contains at most 2.00 wt.% antimony (Sb), - contains at most 7.0 wt.% iron (Fe), and - contains at most 55.00 wt.% zinc (Zn). 14. The method according to any one of claims 1 to 13, wherein the raw material (1, 2) comprises at least one return material from the processing of the molten liquid metal phase and/or the liquid slag phase formed by this method. 15. The method of claim 14, wherein the at least one return material comprises at least one material selected from rejected anodes or other products containing copper, tin and/or lead, an oxide or sulphide of a metal, preferably copper, nickel, tin, lead and/or zinc, including dross containing the metal in the form of an oxide or sulphide formed and removed in a subsequent processing step, a metal silicide, preferably a silicide of a metal selected from copper, zinc, nickel, iron, lead and tin, as well as a crust or other solid substance formed on the wall of a crucible or ladle that was used to move the molten metal and/or molten slag removed from the furnace. 16. The method according to any one of claims 1-15, wherein the raw material (1, 2) is introduced into the center of the liquid bath in the furnace. 17. The method according to any one of claims 1 to 16, wherein the raw material (1, 2) comprises a coarse-grained portion (1), and the average composition of the coarse-grained portion of the raw material fed to the entire batch of loading for melting in the furnace corresponds to at least one, and preferably all, of the following conditions after heating to 1150°C: - contains at least 20% by weight of the total amount of metals, preferably at least 20% of the total amount of copper, nickel, tin, lead and zinc, - contains at least 10.0 wt.% and at most 70.0 wt.% copper (Cu), - contains at least 0.50 wt.% and at most 2.00 wt.% nickel (Ni), - contains at least 1.00 wt.% and at most 8.00 wt.% lead (Pb), - contains at least 0.50 wt.% and at most 2.50 wt.% tin (Sn), - contains no more than 0.10 wt.% antimony (Sb), - contains at least 10.0 wt.% and at most 35.00 wt.% iron (Fe), and - contains at least 2.00 wt.% and at most 15.00 wt.% zinc (Zn). 18. The method according to any one of claims 1 to 17, wherein the content of iron and metals or compounds having at least the same affinity for oxygen as iron or zinc dissolved in the molten metal in the furnace is maintained at a level of at least 1.0 wt.%, preferably at least 1.5 wt.%, wherein the concentration of metals and compounds having at least the same affinity for oxygen as iron or zinc is converted into an equivalent concentration of iron, wherein the equivalent concentration of iron is a concentration of iron that is capable of contributing the same amount of heat of reaction in the form of a metal or compound having at least the same affinity for oxygen as iron or zinc, reacting with oxygen under the conditions of the furnace. 19. The method according to any one of claims 1 to 18, in which the content of iron and metals and compounds having at least the same affinity for oxygen as iron or zinc under furnace conditions, dissolved in the molten metal in the furnace, is maintained at a level of no more than 10.0 wt.%, preferably no more than 9.0 wt.%, wherein the concentration of metals and compounds having at least the same affinity for oxygen as iron or zinc is converted into an equivalent concentration of iron, wherein the equivalent concentration of iron is a concentration of iron that is capable of contributing the same amount of heat of reaction in the form of a metal or compound having at least the same affinity for oxygen as iron or zinc, reacting with oxygen under furnace conditions. 20. The method according to any one of claims 1 to 19, wherein elemental iron is introduced into the smelting stage (100) at a rate at which an excess of iron is maintained in the smelting bath during the method, exceeding its solubility in the metal bath under furnace conditions. 21. The method of claim 20, wherein the amount of excess iron present in the furnace is maintained by at least periodically taking samples of the molten metal phase in the furnace and analyzing them for the presence of iron. 22. The method according to any one of claims 1 to 21, wherein the combustible source of carbon and/or hydrogen is selected from the group consisting of coke, charcoal, soot, hydrocarbon, natural gas, methane, ethane, propane, butane, hydrocarbon which is liquid under normal conditions, hydrocarbon-containing polymer, plastic, plastic waste, fat, oil, paint, varnish, rubber, preferably their waste, as well as combinations thereof. 23. The method according to claim 22, wherein the amount of combustible carbon and/or hydrogen source is maintained below a level at which slag foaming is capable of impairing the performance of the smelting stage, preferably significantly below this level so that the risk of slag foaming remains acceptably low. 24. The method according to any one of claims 1 to 23, wherein at least part of the oxygen-containing gas is introduced close to the interface between the metallic phase and the slag phase. 25. The method according to any one of claims 1 to 24, wherein at least part of the oxygen-containing gas is introduced by means of at least one metal tuyere, the tip of which is immersed in the liquid phase of the slag. 26. The method of claim 25, wherein the gas introduced through the metal tuyere contains at least 30 vol.% oxygen, and more preferably this gas is high purity oxygen. 27. The method of any one of claims 25 or 26, wherein the flow of gas through the metal tuyere provides sufficient cooling to prevent corrosion and/or melting of the tuyere. 28. The method according to any one of claims 1 to 27, wherein at least part of the oxygen-containing gas is introduced into the lower part of the furnace through at least one tuyeres, preferably through a plurality of tuyeres, more preferably a plurality of tuyeres uniformly distributed along the bottom of the furnace. 29. The method according to claim 28, wherein the gas introduced through at least one tuyere is an oxygen-containing gas containing no more than 50 vol.% oxygen, preferably air. 30. The method according to any one of claims 28 or 29, wherein the gas introduced through at least one tuyere is cooler than the molten liquid metal phase surrounding the tuyere. 31. The method according to any one of claims 1 to 30, wherein the smelting slag (5) produced by the method contains at least 20 wt.% iron (Fe). 32. The method according to any one of claims 1 to 31, wherein the composition of the produced smelting slag (5) meets at least one and preferably all of the following conditions: - contains no more than 1.00 wt.% copper (Cu), - contains no more than 0.20 wt.% nickel (Ni), - contains no more than 2.00 wt.% lead (Pb), - contains at most 1.00 wt.% tin (Sn), and - contains at most 22.50% by weight of zinc (Zn). 33. The method according to any one of claims 1 to 32, wherein the composition of said concentrated intermediate copper product (3), which is the main product of the smelting stage (100), meets at least one, and preferably all, of the following conditions: - contains at least 50.0 wt.% copper (Cu), - contains at least 0.01 wt.% nickel (Ni), - contains at least 0.10 wt.% lead (Pb), - contains at least 1.00 wt.% tin (Sn), - contains at least 0.05 wt.% iron (Fe), - contains at least 0.10 wt.% zinc (Zn), and - contains no more than 5% by weight of sulfur (S). 34. The method according to any one of claims 1-33, wherein the liquid bath in the furnace has a temperature in the range of 1100-1300°C. 35. The method according to any one of claims 1 to 34, wherein the exhaust gases from the furnace are collected and treated by cooling and/or filtration. 36. The method according to claim 35, wherein also secondary exhaust gases from the entire furnace are collected and treated by filtration, optionally in combination with cooling. 37. The method according to any one of paragraphs. 1-36, in which the melting (100) is carried out in a melting furnace. 38. The method according to any one of claims 1 to 37, wherein the walls of the furnace are cooled at least partially at their surface. 39. The method according to any one of claims 1 to 38, further comprising the step (400) of fuming the slag phase formed in the smelting step (100) to obtain an evaporated slag (13), wherein the fuming is preferably performed in a fuming furnace. 40. The method according to any one of claims 1-39, further comprising a step (200) of purifying the concentrated intermediate copper product (3) to obtain a purified copper product (8) together with at least one copper refining slag (6). 41. The method according to claim 40, wherein the concentrated intermediate copper product (3) further comprises tin and lead, and which further comprises extracting a metallic composition (7) of raw solder from the concentrated intermediate copper product (3). 42. The method according to claim 41, further comprising the step (500) of extracting from the metallic composition (7) of raw solder at least one of the product (10) of purified soft lead, the product (11) of purified hard lead and the product (12) of purified tin.

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