Sulfur injection in lead slags

Abstract

The invention concerns the recovery of metals from Pb-bearing slags. This goal is achieved by a process for the removal of Pb and Cu from a metallurgical slag, comprising the steps of: providing a smelting furnace; providing a metallurgical slag containing more than 20% by weight of Pb, and further containing Cu, Fe and Zn; heating the slag in the furnace, thereby obtaining a molten slag; providing an amount of elemental sulfur corresponding to at least 110% of the stoichiometric amount to convert 100% of Pb and Cu, and 50% of Fe and Zn to sulfides; feeding said amount of elemental sulfur to the molten slag, thereby obtaining a molten bath comprising a slag phase depleted in Pb and Cu and a matte phase, SO2-bearing off-gas, and flue dust; and, separating the matte phase from the slag phase. This process offers an alternative scheme for the separation and recovery of metals, in particular Pb and Cu, from slags by converting them into metal sulfides. Substituting carbon with sulfur results in the decarbonization of the industrial recovery process.

Description

The field of the present invention is the recovery of metals from Pb-bearing metallurgical slags. In the extraction of metals from ores or from secondary materials containing Cu and Pb, a first step is often a smelting operation producing a Cu matte and a Pb-bearing slag. While a major part of the valuable metals reports to the Cu matte, an appreciable part of them also reports to the slag. Such a slag may contain 20% by weight or more of Pb, together with metals such as Cu, Fe, and Zn. Further treatment is required to recover these valuable metals, especially Cu and Zn. To this end, the slag is typically subjected to reducing conditions in a blast furnace or in an electric arc furnace to produce a metallic alloy containing the valuable metals, and a cleaned slag. The required reducing conditions are obtained by adding carbon or carbohydrates to the furnace, in the form of cokes, or as a liquid or gaseous fuel. The redox reactions lead to the formation of CO, and ultimately to CO2. It is particularly difficult to capture CO2 from the furnace off-gas as the technology is still under development for metallurgical processes. CO2 therefore ends up in the atmosphere, which is undesirable in view of its recognized greenhouse effect. Such a process is for example described in Vanparys et al. (PbZn 2020: 9th International Symposium on Lead and Zinc Processing. The Minerals, Metals & Materials Series. Springer, Cham.). A possible alternative to the addition of carbon or of carbohydrates is based on the use of sulfur (S). The addition of sulfur to metal-containing oxidic phases, such as slags, has indeed been described. Yuan et al. (Powder Technology 230, 63-66, 2012) teaches the addition of sulfur to PbO. Intense cogrinding of the mixture at low temperature results in the partial conversion of Pb to PbS. The sulfides can be separated using flotation. The process may be applied to Pb-bearing wastes such as glasses, wherein Pb is present as an oxide. Han et al. (Selective Sulfidation of Lead Smelter Slag with Sulfur; Metal Mater Trans B 47, 344-354, 2016) shows sulfidation roasting of slags containing 0.5 to 5 %Pb, some Zn and Fe. Sulfur is used instead of more conventional pyrite to avoid Fe as an impurity. Carbon is added to suppress the generation of SO2. This results in the formation of CO and CO2. The initial roasting temperature is limited to about 400 °C to remain below the boiling point of sulfur. A final roasting temperature of up to 1200 °C is applied, while the bulk of the material remains solid. The higher roasting temperatures benefit crystal growth, and result in a more efficient separation of ZnS by flotation in a subsequent process step. CN116024438A describes the introduction of sulfur-containing vulcanizing agents in the form of gypsum, pyrite, or nickel sulfide ore to a molten Ni-rich slag to obtain, among others, a low nickel matte. It appears that most known pyrometallurgical processes make use of carbon or of carbohydrates to obtain the reducing conditions leading to clean slags, depleted in heavy metals. This unavoidably also results in the formation of CO2. It is an objective of the present invention to offer an alternative scheme for the separation and recovery of metals, in particular Pb and Cu, from slags by converting them into metal sulfides. Substituting carbon with sulfur results in the formation of SO2 instead of CO2. Compared to CO2, SO2 is relatively easy to capture. It can be converted to SO3, and then to sulfuric acid using a standard industrial process. Sulfuric acid is a widely used reactant, for example for manufacturing fertilizers. SO2 can also be converted to sulfur when the SO2 concentration in the off-gas is sufficiently high. Environmentally friendly processes, for example using the CaS - SO2 reaction (Environ. Sci. Technol. 2002, 36,13, 3020-3024), have been developed to this end. In this way, a direct effect of the present scheme is the decarbonization of the industrial recovery process. The new process results in the formation of a so-called matte, i.e. a sulfidic phase formed by metal sulfides, instead of the metallic phase, obtained when a traditional carbon-based reduction scheme is applied. A first embodiment describes a process for the removal of Pb and Cu from a metallurgical slag, comprising the steps of: - providing a smelting furnace; - providing a metallurgical slag containing more than 20% by weight of Pb, and further containing Cu, Fe and Zn; - heating the slag in the furnace, thereby obtaining a molten slag; - providing an amount of elemental sulfur corresponding to at least 110% of the stoichiometric amount to convert 100% of Pb and Cu, and 50% of Fe and Zn to sulfides; - feeding said amount of elemental sulfur to the molten slag, thereby obtaining a molten bath comprising a slag phase depleted in Pb and Cu and a matte phase, SO2-bearing off-gas, and flue dust; and, - separating the matte phase from the slag phase. By elemental sulfur is meant sulfur not combined with any chemical element other than sulfur itself. This definition includes octasulfur (Ss), which is a major industrial form of elemental sulfur. This means that for example sulfide concentrates, copper sulfide, iron sulfide, zinc sulfide or sulfur dioxide, are not considered as elemental sulfur according to the present invention. The metallurgical slags typically originate from industrial smelting and refining operations. The most prominent metal in such slags is Pb. Amounts of between 20 and 40% are common. Therefore, in another embodiment, the metallurgical slag contains at most 40% of Pb by weight. Amounts of more than 40% of Pb in such metallurgical slags are less common, as this would make the upstream processes less efficient. Amounts of less than 20% would make the slag viscosity too high and thus require additional fluxing agents. The metallurgical slags according to the present invention typically originate from other industrial processes, in which the majority of the valuable copper has already been removed. Therefore, in another embodiment, the metallurgical slag contains 1 to 5% of Cu by weight. The upper limit of 5% characterizes "copper depleted" slags in contrast to other "copper rich" industrial slags, for which also other methods of refining may be more appropriate. Even dedicated processes for copper removal will typically leave some residual copper in the slag. Having at least 1% of Cu makes a dedicated step for isolating the copper economically worthwhile. In addition to Pb and Cu, other metals, such as Fe, Ni, As, Bi, Sb, Sn and Zn can be present in the metallurgical slag. Therefore, in another embodiment, the metallurgical slag further contains one or more of Ni, As, Bi, Fe, Sb, Sn and Zn. Their respective percentages by weight are typically much lower than the above-mentioned amounts for Pb. All except Fe and Zn are considered as minor impurities with minor or no effect on the present process. The metallurgical slag used at the start of the process can be referred to as "starting slag". Fe will typically be present in an amount of at least 4% and at most 20% by weight in the starting slag. More common are amounts of 4 to 15%, or even 4 to 10%. Generally, Fe is seen as being uncritical and can report to the slag phase, typically in any of its oxide forms, or to the matte phase, typically as FeS. Nevertheless, amounts of more than 20% of iron are less preferred, as this would make the upstream recovery processes less efficient. Zn will typically be present in an amount of at least 2% and at most 15% by weight in the starting slag. More common are amounts of 2 to 10%, or even 2 to 7%. High amounts of Zn are less desired, as Zn not only forms sulfides, but can also fume from the slag and report to the flue dust where it can lead to unwanted downstream condensation. Fe and Zn, if present in significant amounts, should be taken into account when estimating the amount of elemental sulfur to be added to the starting slag. Typically, under the chosen conditions, there is no full conversion to their respective sulfides. One reason forthat is that Fe and Zn have a higher chemical affinity for oxygen than for example Cu and Pb. Therefore, at the present process conditions of temperature and oxygen potential (pO2), the Fe and Zn, present in the initial metallurgical slag, will not convert fully to their respective metal sulfides, but they will remain partially in the depleted slag as metal oxides. It has been observed that an addition of sulfur corresponding to about 50% of the calculated stoichiometric amount for these two metals is sufficient in the present process. Key of the process is the reaction of metals, in particular Pb, in the starting slag with elemental sulfur, thereby forming sulfides. For the most important metals the associated stoichiometric sulfidation reactions are: 2 PbO + 3 S 2 PbS + SO2 2 CU2O + 35^2 CU2S + SO2 2 ZnO + 3 S 2 ZnS + SO2 2 Fe2O3 + 7 S 4 FeS + 3 SO2 At the process conditions, iron oxide and zinc oxide, present in the slag, are typically more stable than lead oxide and copper oxide. As explained above, it can be assumed that in the present process the sulfidation reactions of Zn and Fe proceed to about 50% of completeness. The stoichiometric amount of sulfur (Xs) required for the sulfidation of a metallurgical slag containing Pb, Cu, Zn and Fe, using the above formulae and assumptions for Zn and Fe, can be calculated using following equation: Xs = 3 / 2 Xpb + 3 / 4 Xcu + 0.5*(3 / 2 XZn + 7 / 4 XFe) = 1.5XPb + 0.75Xcu + 0.75XZn + 0.875XFe (with Xm the molar amount of the respective metal M, i.e. Pb, Cu, Zn, Fe) While As, Bi, Ni, Sb and / or Sn can also consume sulfur, they are considered as minor impurities in the present type of slag with very limited or no impact on the overall required stoichiometric amount for the process. In addition to the above-mentioned metals, metallurgical slags also contain slag formers. Typical slag formers are silica (SiO2), alumina (AI2O3), calcium oxide (CaO) or magnesium oxide (MgO). In the metallurgical slags of the present invention, such slag formers can typically be found in the following ranges: 10 to 50 % by weight of silica; 0 to 10 % by weight of alumina; 5 to 60 % by weight of calcium oxide; and, 0 to 10 % by weight of magnesium oxide. Metallurgical slags according to the present invention may contain a combination of slag formers adding up to 40 to 60% by weight. A slag containing a minimum of 40% of slag formers allows for a proper reaction with sulfur, and for an easy phase separation between slag phase and matte phase. Lowering the amount of slag formers to less than 40%, or even to less than 30%, would result in overly concentrated metals, while increasing the amount of slag formers to more than 60%, or even to more than 70%, would result in a diluted overall composition, rendering the recovery of the metals uneconomical. Metallurgical slags generally contain a rather complex mixture of slag formers and metals, typically in form of oxides. Moreover, other impurities can be present, such as P2O5, Na2O or BaO, largely depending on differences in the feed materials that have been used to produce the slag, and / or the applied process conditions. The amounts of the different metals in the Pb-bearing slag are determined upfront by known analytical methods. Based on this, the required stoichiometric amount of sulfur can readily be derived from the above sulfidation reactions. The calculated stoichiometric amount of sulfur is determined by the stoichiometric amount to convert 100% of Pb and Cu, and 50% of Fe and Zn to sulfides. The kinetics of the sulfur reacting with metal oxides in the molten starting slag are limited. The kinetics depend, among others, on temperature and the intimacy of the mixing between the introduced sulfur and the metal oxides in the molten slag. Some sulfur will leave the molten slag unreacted, forming SO2 with any air present above the molten bath or further downstream in the gas cleaning circuit. To compensate for these losses and achieve the desired result of the present process, an excess of S is needed. It is thus required to add more than 100% of the calculated stoichiometric amount of sulfur, such as 110% or more, 120% or more, 130% or more, or 140% or more. In any one of these embodiments, it is preferred to limit the amount of sulfur to at most 300%, preferably to at most 250%, more preferably to at most 200%, and most preferably to at most 150% of the calculated stoichiometric amount of sulfur. In industrial practice, limiting the excess of sulfur fed to the slag is advised due to the cost of the sulfur itself and also to avoid larger off-gas streams. In another embodiment a preferred range for the amount of sulfur is therefore 120 to 150% of the calculated stoichiometric amount of sulfur. Separating the matte phase from the slag phase is typically achieved by selective tapping, when both phases are still liquid. Alternatively, it is possible to tap both phases together, let them cool down and solidify, and separate them either mechanically or by flotation. By "major part" of an element is meant at least 50% by weight of the quantity of that element entering the process. In another embodiment, the furnace is an electric furnace. With electricity as a heat source, the need for carbon-based fuel is alleviated or eliminated. Assuming access to green electricity, the carbon footprint can thus remain remarkably low. In another embodiment, the furnace is a bath furnace or horizontal converter. This type of furnace lends itself well to the injection of sulfur, optionally using a carrier gas. Generally, the partial pressure of oxygen, pO2, in a pyrometallurgical system affects the oxidation state and thermodynamic stability of the metal, slag and gas phases. In this process, the pO2 is preferably in the range of 10'1410 iq-8 It is preferred to avoid the injection of large amounts of oxygen-bearing gasses, such as air, into the molten bath as this will result in sulfur reacting directly with oxygen instead of with the metal oxides in the slag. Moreover, this will produce larger amounts of off-gas to deal with. However, a limited amount of oxygen can be useful if combined with an extra excess of sulfur, thus an additional amount beyond the calculated and required excess according to the above-mentioned stoichiometry for the conversion of metals to sulfides. This additional excess of sulfur can be used as a heat source. Non-reactive gasses such as nitrogen can be injected into the slag. Their effect is that the more volatile species, such as PbS and Zn, will tend to report to the flue dust instead of to the matte phase. This effect can be desired or undesired according to the downstream refining processes available to recover these elements from the flue dust. To accumulate more Pb in the flue dust, a gas stream, for example of nitrogen gas, of 0.5 to 5 Nm^ per kg of Pb is applied. Therefore, in another embodiment, a non-reactive gas is injected into the molten bath during or after the step of feeding elemental sulfur to the molten slag. Thereby, the amount of Pb in the flue dust is increased. Depending on the flow rate, even a majority of Pb can be accumulated in the flue dust. Any amount of Pb reporting to the flue dust decreases the amount of Pb which can report to the matte phase. While other process conditions, such as amount of sulfur, amount of slag formers or temperature, remain the same, this alternative process step allows the skilled person not only to concentrate the Pb in the flue dust, but at the same time separating it from non-volatile other metals, such as copper or iron, which remain in the slag and / or matte phase. Low operational temperatures around the melting point of a slag are preferred to limit refractory degradation in a furnace as much as possible. Preferably, operational temperatures are selected to avoid overheating of the slag too far above its melting point. Typically, the operation temperature of the furnace is from 1150 to 1300 °C, preferably from 1180 to 1230 °C. In another embodiment, the elemental sulfur fed to the molten slag is in liquid form. An advantage of using liquid sulfur is the established technology for its manipulation and transport in an industrial environment. The liquid sulfur can be injected directly into the molten slag, having a positive impact on the process kinetics. In another embodiment, the elemental sulfur is fed by injection through one or more lances or tuyeres. A lance is understood to be a tubular entity dipped from above into the molten bath. A tuyere is understood to be a tubular entity piercing the wall or bottom of the furnace, below or above the level of the molten bath, and equipped with means for the injection of gasses, liquids or solids into the molten bath. A disadvantage of adding solid elemental sulfur, e.g. in form of granules, is the generation of sulfur dust upon manipulation. Sulfur dust shows a potential explosion hazard. Moreover, it will lead to operational difficulties, like clogging of the transportation system. A possible solution for these problems is the use of solid pellets or briquettes comprising sulfur. This has the advantage that sulfur can be top-fed into the furnace, along with other feed materials or fluxes while avoiding the above-mentioned dust generation. Therefore, in another embodiment, the sulfur fed to the molten slag is in the form of solid pellets or briquettes comprising elemental sulfur. In a preferred embodiment, elemental sulfur is mixed with solid metallurgical slag of the same or similar composition as in the first embodiment, then pressed. Pressing increases the relative density of the pellets and improves the mixing of sulfur and metals in the molten bath of the furnace. This enhances the overall reactivity of the sulfur with those metals, making the process more efficient. Therefore, in another embodiment, the solid pellets comprising sulfur further comprise solid slag. In another embodiment, the solid slag used for making pellets has the same composition as the metallurgical slag according to the first embodiment. Variations in the sulfur to slag ratio in this mixture have been tested, namely 4:1, 3:1, 2:1, 1:1, and 1:2. A ratio of 2:1 is preferred. Therefore, in another embodiment, the sulfur to slag ratio in said pellets is from 4:1 to 1:2, preferably from 3:1 to 1:1, more preferably from 2.5:1 to 1.5:1. In another embodiment, more than 80% of the Pb reports to the matte phase as PbS. Pb remaining in the slag after addition of sulfur may result from insufficient addition of sulfur, ineffective reaction behavior between Pb in the molten slag and sulfur, unfavorable reaction kinetics and / or insufficient decantation in the step of separating the matte from the slag. The latter would result in Pb-containing matte droplets in the final / depleted slag and thus an overall higher than expected Pb-content in the slag, despite successful conversion to PbS. In another embodiment, the slag phase depleted in Pb and Cu contains less than 5% of Pb by weight, preferably less than 3%, and most preferably less than 2%. It is technically possible to clean the slag further and reach much lower limits of Pb in the slag. However, this requires additional refining operations, which is currently not economical. In another embodiment, the slag phase depleted in Pb- and Cu contains less than 0.5% of Cu by weight, preferably less than 0.25 %. It has been observed that Cu tends to react faster with the elemental sulfur than Pb. Therefore, very low concentrations of Cu have been observed in the final / depleted slag. Cu reports almost quantitatively to the matte phase as copper sulfide in the present process. In another embodiment, no carbon-based fuel or carbon-based reducing agent is added to the furnace. It is preferred to avoid the addition of carbon-based fuel or carbon-based reducing agent to the furnace. This would result in the formation of CO2, which is undesired when aiming at the decarbonization of the industrial recovery process. Moreover, CO2 is not only more difficult to capture than SO2, it would also make the recovery and re-use of SO2 more difficult due to diluting the latter. In another embodiment, the present process is further comprising a step of converting the SO2 from the S02-bearing off-gas to H2SO4. Modern industrial plants are often equipped with a conversion unit for the production of sulfuric acid. The H2SO4 can for example be valorized or used in acidic leaching operations. In another embodiment, the present process is further comprising a step of converting the SO2 from the S02-bearing off-gas to elemental sulfur. An alternative to the production of sulfuric acid is the conversion of the S02-bearing off-gas to elemental sulfur. This is particularly attractive in combination with an electric furnace, as the off-gas is not diluted by huge amounts of carbon-bearing off-gas, as it would be the case in furnaces burning fuel or cokes. The higher concentration of SO2 in the off-gas of the present process facilitates this alternative conversion method. As an additional advantage, the produced elemental sulfur can be re-used in the present process. Therefore, in another embodiment, the present process is further comprising a step of recirculating the elemental sulfur to the step of feeding elemental sulfur to the molten slag. In another embodiment, the present process is further comprising a step of recovering metals from the matte phase and / or from the flue dust. Downstream refining processes of matte are known and can be combined with the present scheme. One possibility for refining such a matte is to subject it to a converting process. This will produce a metallic phase, which then can be processed in the traditional way, and SO2, which can be captured and / or processed as explained above. Other refining schemes are possible, e.g. using leaching or flotation. Also SO2 originating from a further refining of the obtained PbS and Cu2S-containing matte can be recovered and converted to either H2SO4 or elemental S in one of the above-mentioned conversion operations. The following examples illustrate the invention. Example 1 2625 g of starting slag with composition given in Table 1 was provided. The required stoichiometric amount of elemental sulfur for the sulfidation of this starting slag equals 420 g. 125% of the required stoichiometric amount of sulfur was added to the starting slag, i.e. 525 g. Pellets containing both starting slag and elemental sulfur were prepared. The ratio sulfur to slag in this mixture was 2 to 1. To this end, 270 g of the starting slag was milled and sieved to below 2 mm. The starting slag was mixed with the elemental sulfur to prepare the briquetting mixture. From this mixture pellets were pressed. A 1.5 I alumina crucible with 2355 g of starting slag was placed in an induction furnace and heated under Argon atmosphere at 650 °C / h to a temperature of 1250 °C. Then, the sulfur-slag pellets were added in the course of 2 hours in regular intervals. At the end of the process, 1260 g of final slag and 1295 g of matte were present. The composition of both phases is given in Table 1. Table 1: Mass and composition of slags and matte Mass Pb Cu Fe Zn AI2O3 CaO SiO2 s g % % % % % % % % Starting slag 2625 30.6 4.6 8.9 5.9 3.8 9.3 17.0 - Final slag 1260 3.2 0.1 10.0 6.2 9.1 18.2 35.1 - Matte 1295 57.0 9.7 7.9 6.3 - - - 18.5 Conclusion: The example shows that after addition of 125% of the required stoichiometric amount of sulfur, the Pb content of the final slag was only 3.2%. Thus, 5.0% of the Pb present in the starting slag remained in the slag phase, while 91.9% was transferred to the matte phase. The remainder of the initial amount of Pb, i.e. 3.1%, had ended up in the gas phase and was recovered and analyzed from the flue dust. Example 2 The total amount of sulfur added to the starting slag was doubled with respect to Example 1. 2625 g of starting slag with composition given in Table 2 was provided. The required stoichiometric amount of elemental sulfur for the sulfidation of this starting slag equals 420 g. In this example 250% of the required stoichiometric amount of sulfur was added to the starting slag, i.e. 1050 g. Pellets containing both starting slag and sulfur were prepared. The ratio sulfur to slag in this mixture was 2 to 1. 525 g of starting slag was milled and sieved to below 2 mm. The starting slag was mixed with 1050 g of elemental sulfur to prepare the briquetting mixture. From this mixture pellets were pressed. A 1.5 I alumina crucible with 2100 g of starting slag was placed in an induction furnace and heated under an Argon atmosphere at 650 °C / h to a temperature of 1250 °C. Then, the sulfur-slag pellets were added in the course of 4 hours in regular intervals. At the end of the process, 1150 g of final slag and 1435 g of matte were present. The composition of these phases is given in Table 2. Table 2: Mass and composition of slags and matte Mass Pb Cu Fe Zn AI2O3 CaO SiO2 s g % % % % % % % % Starting slag 2625 30.6 4.6 8.9 5.9 3.8 9.3 17.0 - Final slag 1150 1.5 0.0 7.2 3.9 8.5 22.2 37.2 - Matte 1435 51.8 8.6 10.7 8.0 - - - 20.4 Conclusion: The example shows that after addition of 250% of the required stoichiometric amount of sulfur, the Pb content of the final slag was only 1.5%. Thus, 2.1% of the Pb present in the starting slag remained in the slag phase, while 92.5% was transferred to the matte phase. Just like in Example 1, the remainder of the initial amount of Pb, i.e. 5.3%, had ended up in the gas phase and was recovered and analyzed from the flue dust. Example 3 (Comparative Example) 2625 g of starting slag with composition given in Table 3 was provided. The required stoichiometric amount of elemental sulfur for the sulfidation of this starting slag equals 420 g. In this example 62% of the required stoichiometric amount of sulfur was added to the starting slag, i.e. 260 g. Pellets containing both starting slag and sulfur were prepared. The ratio sulfur to slag in this mixture was 2 to 1. 130 g of starting slag was milled and sieved to below 2 mm. The starting slag was mixed with 260 g of elemental sulfur to prepare the briquetting mixture. From this mixture pellets were pressed. A 1.5 I alumina crucible with 2495 g of starting slag was placed in an induction furnace and heated under an Argon atmosphere at 650 °C / h to a temperature of 1250 °C. Then, the sulfur-slag pellets were added in the course of 1 hour in regular intervals. At the end of the process, 1805 g of final slag and 890 g of matte were present. The composition of these phases is given in Table 3. Table 3: Mass and composition of slags and matte Mass Pb Cu Fe Zn AI2O3 CaO SiO2 s g % % % % % % % % Starting slag 2625 30.6 4.6 8.9 5.9 3.8 9.3 17.0 - Final slag 1805 10.6 0.1 11.9 7.4 5.9 13.4 24.5 - Matte 890 66.0 13.9 2.4 2.1 - - - 15.0 Conclusion: The example shows that after addition of 62% of the required stoichiometric amount of sulfur, the Pb content of the final slag was 10.6%. Thus, 23.8% of the Pb present in the starting slag remained in the slag phase, while only 73.2% was transferred to the matte phase. The amount of elemental sulfur added in the reaction was thus not sufficient. The remainder of the initial amount of Pb, i.e. 3.0%, had ended up in the gas phase and was recovered and analyzed from the flue dust.

Claims

1. Process for the removal of Pb and Cu from a metallurgical slag, comprising the steps of:- providing a smelting furnace;- providing a metallurgical slag containing more than 20% by weight of Pb, and further containing Cu, Fe and Zn;- heating the slag in the furnace, thereby obtaining a molten slag;- providing an amount of elemental sulfur corresponding to at least 110% of the stoichiometric amount to convert 100% of Pb and Cu, and 50% of Fe and Zn to sulfides;- feeding said amount of elemental sulfur to the molten slag, thereby obtaining a molten bath comprising a slag phase depleted in Pb and Cu and a matte phase, SO2-bearing offgas, and flue dust; and,- separating the matte phase from the slag phase,wherein the pO2 of the process is in the range of 10-14 to 10-8, and,wherein the operation temperature of the furnace is from 1150 to 1300 °C, preferably from 1180 to 1230 °C.

2. Process accordingto claim 1, wherein the furnace is an electric furnace.

3. Process according to claim 1 or 2, wherein the elemental sulfur fed to the molten slag is in liquid form.

4. Process according to any one of claims 1 to 3, wherein the elemental sulfur is fed by injection through one or more lances or tuyeres.

5. Process according to claim 1 or 2, wherein the elemental sulfur fed to the molten slag is in the form of solid pellets comprising sulfur.

6. Process accordingto claim 5, wherein the solid pellets comprisingsulfurfurther comprise solid slag.A23048-WO-PCT7. Process accordingto claim 6, wherein the solid slag has the same composition as the metallurgical slag accordingto claim 1.

8. Process according to claims 6 or 7, wherein the ratio of sulfur to solid slag is from 4:1 to 1:2.

9. Process according to any one of claims 1 to 8, wherein more than 80% of the Pb reports to the matte phase as PbS.

10. Process according to any one of claims 1 to 8, wherein a non-reactive gas is injected into the molten bath during and / or after the step of feeding elemental sulfur to the molten slag.

11. Process accordingto any one of claims 1 to 10, wherein the slag phase depleted in Pb and Cu contains less than 5% of Pb by weight, preferably less than 3%, and most preferably less than 2%.

12. Process accordingto any one of claims 1 to 11, wherein the slag phase depleted in Pb and Cu contains less than 0.5% of Cu by weight, preferably less than 0.25%.

13. Process accordingto anyone of claims 1 to 12, wherein no carbon-based fuel or carbon-based reducing agent is added to the furnace.

14. Process according to any one of claims 1 to 13, further comprising a step of converting SO2 from the SO2-bearing off-gas to H2SO4.

15. Process according to any one of claims 1 to 13, further comprising a step of converting SO2 from the SO2-bearing off-gas to elemental sulfur.

16. Process accordingto claim 15, further comprising a step of recirculating the elemental sulfurto the step of feeding elemental sulfurto the molten slag.A23048-WO-PCT17. Process according any one of claims 1 to 16, further comprising a step of recovering metals from the matte phase and / or from the flue dust.