WO2011059380A1

WO2011059380A1 - Process for recovering a valuable metal from a sulfidic material by hydrometallurgy

Info

Publication number: WO2011059380A1
Application number: PCT/SE2010/051199
Authority: WO
Inventors: Jan-Eric Sundkvist
Original assignee: Boliden Mineral Ab
Priority date: 2009-11-10
Filing date: 2010-11-03
Publication date: 2011-05-19
Also published as: SE0950848A1; SE534315C2

Abstract

The present disclosure relates to a process for recovering a valuable metal from a sulfidic material. The process comprises subjecting the sulfidic material to a hydrometallugical step followed by selective removal of elemental sulful from the metallurgical residue before a final metal recovery step, such as cyanide leaching. The process improves the over-all process economy and reduces the cyanide consumption as a result of the selective removal of elemental sulfur.

Description

PROCESS FOR RECOVERING A VALUABLE METAL FROM A SULFIDIC MATERIAL BY HYDROMETALLURGY

The present disclosure relates in general to a process for recovering a valuable metal from a sulfidic material comprising said valuable metal by hydrometallurgy, preferably bioleaching followed by cyanide leaching.

Background

Sulfidic materials, such as ores and ore concentrates, can be leached in the presence of oxygen and microorganisms capable of favoring oxidation of both sulfur and iron and other metals in the materials, with the object of winning the valuable metal content of the material. This type of leaching is called bioleaching or biomining and is as such commonly known.

The ores or concentrates usually comprise a mixture of several minerals, such as chalcopyrite, arsenopyrite and iron pyrite, and may also comprise secondary sulfides. The noble and/or precious metal content is often embedded in for example arsenopyrite or iron pyrite, i.e. so called refractory precious metals.

Valuable metals can be leached out by bioleaching and brought into solution. The formed solution can then be treated for selective winning of valuable metals, such as copper, nickel, cobalt etc. The noble or precious metal content that cannot be recovered by leaching in this way can also be recovered by first dissolving any surrounding metal sulfides so as to free the noble or precious metals, and thereafter treating the bioleaching residue hydrometallurgically in a conventional manner to leach out the noble or precious metals.

Bioleaching processes afford certain advantages over other possible hydronnetallurgical processes for working-up metal containing sulfidic materials, for instance pressure leaching, by virtue of the fact that microorganisms will favor oxidation of both sulfidic sulfur and elemental sulfur to form sulfates at low temperatures and atmospheric pressure. Moreover, oxidation of Fe(ll) to Fe(lll) as well as As(lll) to As(V) is also favored. The leached material should therefore at best be possible to leach further in subsequent stages, for instance in precious metal recovery processes, without any risk of problems caused by the presence of elemental sulfur. It has however been found that also minor amounts of elemental sulfur, present in the bioleached material, present problems in a subsequent cyanide leaching process for recovery of precious metals, and recovery of gold and other precious metals from bio-oxidation residues has therefore historically been associated with high cyanide consumption.

There has been a debate through the years regarding the role of the enzyme rhodanese and how it influences the cyanide consumption. Lawson, The role of the enzyme rhodanese in cyanide consumption for gold extraction from bacterially leached refractory ores, IBS'97 Conference proceedings, Biomine International Conference, 4-6 August 1997, discloses that the enzyme rhodanese catalyses formation of thiocyanate.

However, Gardner et al., Production of rhodanese by bacteria present in bio-oxidation plants used to recover gold from arsenopyrite concentrates, Journal of Applied Microbiology 2000, 89, 185-190, discloses that it is not clear whether the rhodanese activity contributes significantly to the loss of cyanide in the gold extraction process or whether this loss is purely a result of chemical reactions. Gardner et al. concludes that it is possible that the loss of cyanide is directly attributed to chemical reactions and only indirectly to biological effects.

Moreover, Aswegen et al., New developments in Bacterial Oxidation Technology to Enhance the Efficiency of the BIOX® Process, Bac-Min

Conference, Benigo, Vic, 8-10 November 2004, discloses that tests have shown that elemental sulfur and other intermediate sulfur oxidation species are the main cyanide consumers.

Thus, it is well known that the elemental sulfur content in a bio-residue from mesophilic or moderate thermophilic bio-oxidation could be the major consumer of cyanide in a cyanidation process by formation of thiocyanate (SCN^").

S° + CN^"→ SCN^"

It is also known that thiocyanate may be formed by cyanolysis of thiosalts formed in grinding, flotation and cyanide leaching processes by oxidation or disproportionation of elemental sulfur and pyrite.

Thus, in order to minimize the consumption of cyanide in a recovery process without impairing the recovery of valuable metals, it is necessary to overcome the problem of elemental sulfur and reduced inorganic sulfur

compounds. In addition to the high direct operating cost for the dissolvable elemental sulfur content of the sulfidic material, the formed thiocyanate, reduced inorganic sulfur compounds and any residual elemental sulfur in the leach residue from the cyanidation process may have an overall negative environmental impact. SCN^" is a rather stable compound, which decomposes slowly to sulfate, carbon

dioxide/carbonate species and ammonium/ammonia under oxidizing conditions. It has also been found that SCN^" will pass through a conventional cyanide destruction process such as the INCO process. To decompose SCN^" rapidly and completely, Caro's acid (H₂SO₅) has to be applied and neutralized, which contributes to a high operating cost.

Any residual SCN^" in effluent from a tailings pond may be a potential acid producer and oxygen consumer. Moreover, since all conventional

cyanide/thiocyanate destruction processes ends up with soluble nitrogen species, minimization of cyanide addition rate is of major interest to reduce the nitrogen load on the recipient.

Furthermore, elemental sulfur in bio-residues is also unwanted since it may capsulate for example fine gold particles and thereby slow down the precious metals leaching rate. Any residual elemental sulfur discharged to tailings ponds or other deposit facilities may also be harmful in the long term due to its instability in water at neutral and alkaline conditions.

Lindstrom et al, A sequential two-set process using moderately and extremely thermophilic cultures for biooxidation of refractory gold concentrates, Hydrometallurgy 71 (2003) 21 -30, discloses a sequential two-step bioleaching process for gold-containing refractory pyrite/arsenopyrite concentrates. A moderately thermophilic culture was used in the first stage and an extremely thermophilic culture, S. metallicus, was used in the second stage. It was inter alia found that the cyanide consumption during cyanidation of the leached mineral residues was significantly reduced by the two-step bioleaching process compared to a process comprising a single bioleaching step using a moderately thermophilic culture. Lindstrom et al. concluded that this was probably due to a lower final concentration of cyanide reactive sulfur compounds after biooxidation with S. metallicus and/or the fact that fewer microorganisms were attached to the mineral residue. Similar results have also been reported in the article by Aswegen et al. mentioned above.

There has been a focus in recent years on research on bioleaching processes for copper extraction from chalcopyrite, which is the mineral that holds the main copper reserves in the world. A major obstacle to an efficient process for bioleaching chalcopyrite is the formation of a barrier layer over the surface of the mineral that hinders the passage of the chemical oxidant. This phenomenon is often called passivation or hindered dissolution. It has recently been discovered that a thick layer of elemental sulfur is initially formed on the surface of the minerals. Leaching can sometimes continue for a while due to peeling of the sulfur layer, even though the leaching rate is generally low. After a while a permanent jarosite layer is formed and leaching is stopped.

In order to achieve an efficient process for leaching chalcopyrite ores or concentrates such that leaching of chalcopyrite can reach its full potential, it is necessary to overcome the problem of passivation.

Klauber, C. 2008. Int J Min Process 86:1 -17, states that elemental sulfur formation remains a systemic phase in the context of heap bioleaching but is not a problem of any consequence for mixed culture systems unlike jarosite, which causes true hindered dissolution problems. The conclusion by Klauber is that there are three ways to go for preventing true passivation, namely to prevent jarosite formation by maintaining a very low pH in the process, cause jarosite to self- precipitate and not to coat the chalcopyrite, and remove iron from the heap.

According to Klauber, iron control by jarosite precipitation would be the preferred route.

Thus, in accordance with conclusions made by Klauber, the problem of passivation or hindered dissolution of chalcopyrite seems only to a minor degree be related to elemental sulfur. This however still remains to be further investigated.

Summary

The object of the invention is to overcome, or at least significantly reduce, the problems associated with elemental sulfur in hydrometallurgical residues subjected to downstream processing for recovery of a valuable metal. The object is achieved by the process for recovering a valuable metal from a sulfidic material comprising said valuable metal in accordance with independent claim 1 . Embodiments of the process are defined by the dependent claims.

In accordance with a first aspect of the invention, the process comprises the steps of subjecting the sulfidic material to at least a first hydrometallurgical step and subjecting the discharged slurry to a solid/liquid separation step thus obtaining a metallurgical residue, subjecting the metallurgical residue to a step for selective removal of elemental sulfur, wherein said step for selective removal of elemental sulfur is conducted at a pH of 3.0 or higher, and subjecting the metallurgical residue to a final metal recovery step after said step for selective removal of elemental sulfur.

The first hydrometallurgical step is preferably a bioleaching step, but may also be any other hydrometallurgical step conventionally used for treatment of sulfidic materials for enabling recovery of valuable metals, such as autoclaving.

It will be readily apparent to the skilled person that also reduced inorganic sulfur compounds may be selectively removed during the step for selective removal of elemental sulfur.

In accordance with a second aspect of the invention, the process for recovering a valuable metal from a sulfidic material comprises the steps of subjecting the sulfidic material to at least a first hydrometallurgical step, for example a bioleaching step, and subjecting the slurry from the hydrometallurgical step to a solid/liquid separation step thus obtaining a metallurgical residue, subjecting the metallurgical residue to a biooxidation step for selective removal of elemental sulfur, wherein said biooxidation step is conducted at a pH of 3.0 or higher, and thereafter subjecting the metallurgical residue to a final metal recovery step. Thus, elemental sulfur will be selectively removed by oxidation during said biooxidation step at a pH of 3.0 or higher. Moreover, also reduced inorganic sulfur compounds will be selectively removed during the biooxidation step.

According to a preferred embodiment, the biooxidation step for selective removal of elemental sulfur is conducted at a pH of 3.5-6, preferably at a pH of 3.8-5.5.

In accordance with a third aspect of the invention, the process for recovering a valuable metal from a sulfidic material comprises the steps of subjecting the sulfidic material to at least a first hydrometallurgical step, for example a bioleaching step, and subjecting the slurry from the hydrometallurgical step to a solid/liquid separation step thus obtaining a metallurgical residue, subjecting the metallurgical residue to a sulfite leaching step for selective removal of elemental sulfur, wherein said sulfite leaching step is conducted at a pH of 3.0 or higher, preferably at a pH of at least 3.5, and thereafter subjecting the metallurgical residue to a final metal recovery step. Thus, elemental sulfur will be selectively removed by dissolution during said sulfite leaching step.

According to a preferred embodiment, the sulfite leaching step is performed at a temperature of above 25 °C, preferably above 40 °C.

According to another preferred embodiment, sulfite is added to the sulfite leaching step in an amount which is over-stoichiometric relative to the elemental sulfur demand. Preferably, sulfite is added in an amount of at least twice the stoichiometric demand of elemental sulfur.

In accordance with yet another preferred embodiment, the metallurgical residue is subjected to a thiosulfate leaching step after said sulfite leaching step. The thiosulfate needed in said thiosulfate leaching step may be generated in situ during the sulfite leaching step.

It is an essential feature of the process according to the invention that the step for selective removal of elemental sulfur is performed at a pH of at least 3.0 since soluble ferric ions, Fe(lll), are not stable in the solution at such a high pH. Thus, the solution will be substantially free from ferric and ferrous ions. When Fe(lll) is present in the solution, it will continue to contribute to the formation of elemental sulfur and ferrous iron from residual sulfidic material, and any process step for selectively removing sulfur would in such a case be insufficient since there would be a continuous formation of elemental sulfur in parallel to the removal. However, at pH values of about 3.0 and above, a major part of Fe(lll) will precipitate and will not affect the solids of the process. Thereby, the elemental sulfur can be selectively removed by for example biooxidation or anaerobic sulfite leaching.

The process according to the present invention thus enables efficient removal of elemental sulfur which in turn leads to a significantly improved over-all process economy. For example, in case the final metal recovery step is a cyanide leaching step, the cyanide consumption is substantially reduced. In fact, it has been found that the cyanide consumption can be reduced by at least 90 % while still achieving the ultimate gold recovery when leaching sulfidic materials comprising gold.

The process according to the invention also results in lower costs for destruction of process solution and lower nitrogen load on the recipient which is desirable from an environmental point of view.

Moreover, the leaching rate is improved. For example, it has been found that the initial leaching rate of gold and silver are significantly enhanced by the removal of elemental sulfur. The retention time during the cyanidation process can also be significantly reduced which also contributes to the lower cyanide consumption.

Furthermore, it has been found that when leaching chalcopyrite

concentrates or ores, the formation of layers of elemental sulfur and/or jarosite which normally occurs on the surfaces of the minerals during leaching can be significantly reduced.

All of the above identified advantages are achieved while still not reducing the amount of valuable metal recovered by the process compared to previously known processes.

Brief description of the drawings

Figure 1 illustrates schematically a process according to one embodiment of the invention wherein the metallurgical residue is subjected to a second bioleaching step conducted at a pH of at least 3.0.

Figure 2a illustrates schematically a process according to another embodiment of the invention wherein the metallurgical residue is subjected to a sulfite leaching step at a pH of 3.5-8.

Figure 2b illustrates schematically a process according to yet another

embodiment of the invention wherein the metallurgical residue is subjected to a sulfite leaching step at a pH of at least 3.0 followed by a bioleaching step.

Figure 3 illustrates schematically a process according to yet another

embodiment of the invention wherein the metallurgical residue is subjected to a sulfite leaching step at a pH of 3.5-8 followed by a thiosulfate leaching step and thereafter a metal recovery step by cementation. Figure 4 illustrates schematically a process according to yet another embodiment of the invention wherein chalcopyrite is leached by heap leaching and subjected to a step for selective removal of elemental sulfur.

Figure 5 shows the sulfur removal profile for a batch-wise biooxidation of a bioresidue at a pH of about 4.5 and about 20 % solids.

Figure 6 shows the sulfur removal profile for a batch-wise biooxidation of a bioresidue at a pH of about 4.5 and about 20 % solids followed by sulfite leaching.

Figure 7 shows test results of the residual sulfur content in a bioresidue

versus the leach time in hours when sulfite has been added in an amount of 1 .1 , 2.3 and 3.5, respectively, times the stoichiometric demand of elemental sulfur for a sulfite leaching at about 45 °C, about pH 8.5 and 18 % solids.

Figure 8 shows test results of the sulfur conversion versus the leach time for the same conditions as in Figure 5.

Figure 9 shows test results of the residual sulfur content in a bioresidue

versus the leach time in hours when sulfite has been added in an amount of 1 .1 and 3.3, respectively, times the stoichiometric demand of elemental sulfur for a sulfite leaching at about 65 °C, about pH 7.5 and 18 % solids.

Figure 10 shows test results of the sulfur conversion versus the leach time for the same conditions as in Figure 9.

Figure 1 1 shows the test results of the gold in leach tails versus the leach time obtained for the same conditions as in Figure 9.

Figure 12a shows test results of the NaCN consumption versus the leach time during cyanide leaching for a bioresidue wherein elemental sulfur has not been removed, a bioresidue wherein elemental sulfur has been removed by bioleaching at a pH or about 3.5, and a bioresidue wherein elemental sulfur has been removed by sulfite leaching.

Figure 12b shows test results of the thiocyanate formation for the same

conditions as in Figure 12a.

Figure 13 shows the obtained leaching profile for a copper concentrate which is bioleached, followed by sulfite leaching and subsequent bioleaching. shows the obtained leaching profile of a copper concentrate which is bioleached, followed by sulfite leaching and subsequent bioleaching, in a repeated test. Detailed description

The invention will be further described below with reference to the accompanying drawings. It should however be noted that the invention is not limited to the embodiments described below and shown in the drawings, but may be modified within the scope of the claims.

The sulfidic material may be any sulfidic material comprising a noble and/or precious metal or other valuable metals and includes any ore or

concentrate which may be hydrometaNurgically treated for recovery of the valuable metal, for example bioleached. Such ores or concentrates usually comprise a mixture of several minerals, as well as secondary sulfides, as previously disclosed. The process is especially suitable for treatment of refractory gold-arsenic ores or concentrates wherein the gold generally is enclosed in arsenopyrite or iron pyrite. The process may also be used for treatment of chalcopyrite.

The process for recovering a valuable metal from a sulfidic material according to the invention comprises subjecting the sulfidic material to at least a first hydrometallurgical step. The hydrometallurgical step is preferably a

bioleaching step, for example a continuous bioleaching process. It is however possible to use other types of hydrometallurgical steps previously known, for example autoclaving.

When the first hydrometallurgical step is a bioleaching step, said

bioleaching step is performed in accordance with previously known processes and is thus conducted at a pH value of less than 3, typically a pH of equal to or less than 2.5, in order to ensure optimal growth and performance of the biomass. Any type of biomass suitable for bioleaching of the sulfidic material may be used, such as mesophilic biomass or thermophilic biomass. The bioleaching step could for example be a continuous bioleaching, wherein the material is leached

continuously in a number of reactors connected in parallel or in series, or a heap leaching step.

Naturally, the process may also comprise subjecting the sulfidic material to a plurality of bioleaching steps or other hydrometallurgical steps in accordance with conventional techniques without departing from the invention. Moreover, the process may also include steps preceding the first hydrometallurgical step, such as a pre-oxidation, if desired. Such steps are also performed in accordance with previously known techniques.

After the sulfidic material has been bioleached or subjected to other hydrometallurgical steps such as autodaving, the obtained slurry is subjected to a solid/liquid separation step and the solution, for example comprising iron and arsenic, is removed for processing in accordance with previously known

techniques, such as neutralization.

It will be readily apparent to the skilled person that when the process is a heap leaching process, for example for treatment of chalcopyrite ores or concentrates, the solid/liquid separation step is a wash/ rinse cycle.

The solids obtained from the solid/liquid separation step, i.e. the

metallurgical residue, are thereafter subjected to a step for selective removal of elemental sulfur, wherein said step is conducted at a pH of 3.0 or higher. The selective removal of elemental sulfur may for example be performed either by biooxidation or by chemical dissolution.

Due to the fact that the step for selective removal of elemental sulfur is performed at said pH, the solution will only comprise very low amounts of ferric and ferrous ions, if any. This is important since ferric ions in the solution act as an oxidizing agent for formation of new elemental sulfur by oxidation of remaining sulfidic material, such as arsenopyrite or pyrite. By minimizing the amount of ferric ions in the solution, the risk of formation of new elemental sulfur for example by aeration in the presence of microorganisms is thus minimized.

Furthermore, the fact that the step for selective removal of elemental sulfur is conducted at a pH of 3.0 or higher, and at a moderate temperature, gives a considerably less corrosive environment compared to other methods for removal of sulfur or sulfur species, such as biooxidation at a pH of about 2.5, and thus reduces the requirement of corrosion resistance of the used construction materials.

After the elemental sulfur has been selectively removed the metallurgical residue is subjected to a final metal recovery step in accordance with previously known techniques. The final metal recovery step can for example be a cyanide leaching step, such as a carbon-in-leach (CIL) process, a carbon-in-pulp (CIP) process, a resin-in-pulp (RIP) process and a Merrill-Crowe process. The final metal recovery step may also be a thiosulfate/NH /NH₃/Cu leaching process or a metal recovery step by cementation.

Naturally, the metallurgical residue may also be subjected to one or more additional steps after said step for selective removal of elemental sulfur but prior to the final metal recovery step. Such additional steps includes, but are not limited to, a solid/liquid separation step, a bioleaching step or a biooxidation step, a sulfite leaching step and a thiosulfate leaching step.

In accordance with one aspect of the invention, the step for selective removal of elemental sulfur is a biooxidation step whereby elemental sulfur is oxidized by the assistance of a suitable biomass. The elemental sulfur is primarily oxidized to SO₄ ^2", but also to a minor extent to S2O3^2" and/or S_nO6^2". The solution comprising these sulfur containing species can be removed in a conventional solid/liquid separation step, recycled and used in a preceding bioleaching step if desired. It is however also possible to omit such a solid/liquid separation step and transfer the slurry from the biooxidation step for selective removal of elemental sulfur directly to a final metal recovery step, such as a cyanide leaching step.

The maximum growth rate of biomass might be somewhat slower during said biooxidation step for selective removal of elemental sulfur compared to a conventional bioleaching step due to the higher pH. In order to compensate for the slower growth rate of biomass, the biooxidation step could preferably be

conducted for a longer period of time than if it was conducted at a conventional pH, such as equal to or below 2.5. It has been shown that the relatively limited generation of heat from selective removal of typical residual elemental sulfur contents may require a longer retention time than the critical for wash-out in order maintain a suitable operating process temperature, which depends on the actual elemental sulfur concentration and pulp solid content.

The upper limit of the pH which may be used during the biooxidation step for selective removal of elemental sulfur depends on the type of biomass used. However, it is preferred that the pH is kept below 6 in order to achieve an efficient process since most types of suitable microorganisms have a very low growth rate above pH 6. Thus, in accordance with one preferred embodiment of the process, the biooxidation step for selective removal of elemental sulfur is conducted at a pH of 3.5-6, more preferably at a pH of 3.8-5.5. Moreover, the biooxidation step should naturally be performed without limitation of oxygen and essential nutrients in order to ensure sufficient growth and activity of the biomass. The oxygen is supplied either by air, pure oxygen or oxygen enriched air. The pH is preferably controlled by limestone addition which also will serve as a carbon dioxide source for growth. Furthermore, it will be readily apparent to the skilled person that the temperature during the biooxidation step is adapted to the biomass primarily inoculated or developed by time as well as to the actual heat balance.

In accordance with another aspect of the invention, the step for selective removal of elemental sulfur is a sulfite leaching step, preferably conducted at a pH of at least 3.5. Sulfite species dissolve elemental sulfur in accordance with the following formulas:

S° + SO₃ ^2"→ S₂O₃ ^2"

or

The sulfite leaching should be performed in the absence of air in order not to negatively influence the leaching process by unwanted oxidation of the remaining sulfidic material and added sulfite ions. Thus, the sulfite leaching step is an anaerobic leaching step.

It has been found that the dissolution rate of sulfur during the sulfite leaching step depends on the temperature. Therefore, the sulfite leaching should preferably be performed at an increased temperature in order to achieve an efficient process. Thus, the sulfite leaching step is preferably conducted at a temperature above 25 °C, more preferably at a temperature above 40 °C. In accordance with one preferred embodiment, the sulfite leaching step is conducted at a temperature of at least 50 °C.

Furthermore, it has been found that the sulfur extraction rate depends on the total sulfite addition. Sulfite should preferably be added in an over

stoichiometric amount in relation to the elemental sulfur content. In accordance with a preferred embodiment, sulfite is added in an amount of at least twice the stoichiometric demand relative to the elemental sulfur content. More preferably, sulfite is added in an amount of at least 2.5 times the stoichiometric demand.

In recent years, cyanide leaching has been banned in many regions for environmental reasons. Thiosulfate, in combination with ammonia/ammonium and Cu²⁺, is generally considered to be the most promising alternative today to cyanide leaching for extraction of gold. Thus, when the process comprises a thiosulfate leaching step for recovery of precious metals it is advantageous from a process economy point of view if thiosulfate could be generated in-situ and at the same time remove the unwanted residual elemental sulfur. This has been found possible with the process according to the invention, especially if the sulfite leaching is performed at a pH of at least 3.5 since the thiosulfate formed is relatively stable during these conditions.

Thus, in accordance with one embodiment of the invention, the sulfite leaching step should preferably be performed at a pH of at least 3.5 in order to ensure sufficient stability of the formed thiosulfate. It has been found that at sulfur contents normally present in bioleaching residues, the process generates the thiosulfate content which is necessary during leaching of noble metals. The source of sulfite can for example be Na2SO3, NaHSO3, Na2S2O₅ or CaO/SO2.

The obtained leach solution, separated from the sulfite leaching step comprises reduced inorganic sulfur species, mainly thiosulfate. The solution may suitably be recycled back to a preceding bioleaching step as make-up water.

Alternatively, the solution may be diverted to a separated biooxidation reactor for destruction of formed thiosalts.

Figure 1 illustrates schematically one embodiment of the process according to the invention. A sulfidic material, for example a refractory gold- arsenic concentrate, is subjected to a bioleaching step 1 at a pH of about 1 .0-2.0. The slurry resulting from the bioleaching step is transferred to a solid/liquid separation step 2 from which the solution comprising Fe and As is removed to neutralization, as indicated by arrow 3. The solids are subjected to a biooxidation step 4, which is conducted at a pH of at least 3.0 such that the amount of ferric and ferrous ions in the solution is minimized. The redox potential is maintained above the stability domain of elemental sulfur. In the biooxidation step 4, elemental sulfur is thus selectively oxidized at a significantly faster rate than the formation rate of elemental sulfur or other reduced species from the solids. The pulp resulting from the biooxidation step 4, and which comprises SO₄ ^2", as well as minor amounts of S2O3^2" and/or S₄O6^2", is transferred as illustrated by arrow 5 to a cyanide leaching step (CIL) 6 wherein for example the gold is recovered. The cyanide leaching step is typically conducted at a pH of about 10.5. Even though not illustrated in Figure 1 , a sulfite leaching step may be incorporated between the biooxidation step 4 and the cyanide leaching step 6 if desired. Such a sulfite leaching step would remove possible residual elemental sulfur from the biooxidation step 4.

Figure 2a illustrates schematically another embodiment of the process according to the invention. A sulfidic material, for example a refractory gold- arsenic concentrate, is subjected to a bioleaching step 1 at a pH of about 1 .0-2.0. The slurry resulting from the bioleaching step is transferred to a solid/liquid separation step 2 from which the solution comprising Fe and As is removed to neutralization, as indicated by arrow 3. The solids are subjected to a sulfite leaching step 7 at a pH of at least 3.5, preferably a pH of 3.5-8, in order to selectively dissolve elemental sulfur. The pulp from the sulfite leaching step 7, which comprises S2O3^2", is thereafter transferred to a solid/liquid separation step 9 as indicated by arrow 8. The solution from the separation step 9, and which comprises thiosulfate, is recycled back to initial bioleaching step, as illustrated by arrow 10, wherein it is oxidized to sulfate. The solids are transferred to a cyanide leaching step (CIL) 6 wherein for example the gold is recovered. The cyanide leaching step is typically conducted at a pH of about 10.5.

Figure 2b illustrates schematically yet another embodiment of the process according to the invention. A sulfidic material, for example a refractory gold- arsenic concentrate, is subjected to a bioleaching step 1 at a pH of about 1 .0-2.0. The slurry resulting from the bioleaching step is transferred to a solid/liquid separation step 2 from which the solution comprising Fe and As is removed to neutralization, as indicated by arrow 3. The solids are subjected to a sulfite leaching step 7 at a pH of at least 3.5, preferably a pH of 3.5-8, in order to selectively dissolve elemental sulfur. The pulp from the sulfite leaching step 7, and which mainly comprises S2O3^2", is thereafter transferred to a biooxidation step 1 1 which is conducted at a pH of at least 3.0, preferably at a pH of at least 3.5. The pulp is thereafter subjected to a cyanide leaching step (CIL) 6 wherein the gold is recovered. The cyanide leaching step is typically conducted at a pH of about 10.5.

Compared to the embodiment shown in Figure 2a, the embodiment shown in Figure 2b does not require the second solid/liquid separation step 9, which in some cases may be advantageous. However, in the embodiment shown in Figure 2b, an additional biooxidation step is included. The additional biooxidation step 1 1 will however oxidize reduced thiosalts in the sulfite leaching solution, which reduces the cyanide consumption in the subsequent cyanide leaching step as it minimizes the formation of SCN^".

Figure 3 illustrates schematically yet another embodiment of the process according to the invention. A sulfidic material, for example a refractory gold- arsenic concentrate, is subjected to a bioleaching step 1 at a pH of about 1 .0-2.0. The slurry resulting from the bioleaching step is transferred to a solid/liquid separation step 2 from which the solution comprising Fe and As is removed to neutralization, as indicated by arrow 3. The solids are subjected to a sulfite leaching step 7 at a pH of at least 3.5, preferably a pH of 3.5-8, in order to selectively dissolve elemental sulfur.

In recent years, cyanide leaching has been banned in many regions for environmental reasons as mentioned above. Thiosulfate, in combination with ammonia and Cu²⁺, is generally considered to be the most promising alternative to cyanide leaching for extraction of gold. Thus, in accordance with the embodiment shown in Figure 3, the pulp from the sulfite leaching step 7, and which comprises S2O3^2", is thereafter transferred to a thiosulfate leaching step 12.

The thiosulfate leaching step is suitably conducted at a pH of about 8-10 and in the presence of Cu²⁺ and NH 7NH₃. The thiosulfate is produced in-situ, which has a significant beneficial effect on the overall operating cost.

The slurry is thereafter subjected to a solid/liquid separation step 13 and the solids are removed as indicated by arrow 14. The solution, comprising Au, Ag, S2O3^2" and S₄O6^2" is transferred as indicated by arrow 15 to a metal recovery step 16 wherein silver and gold are recovered by cementation with e.g. elemental copper. The solution from the metal recovery step 16 and which comprises S2O3^2" and S₄O6^2" is recycled to the bioleaching step 1 as indicated by arrow 17.

Figure 4 shows an embodiment of the invention wherein a chalcopyrite concentrate or ore is leached by heap leaching. The process comprises a primary stage for copper leaching at a pH of below 3. The pregnant solution is transferred to a pregnant solution pond and thereafter to a solvent extraction-electrowinning process SX/EW, and the acid is reused in the primary stage. The suitable leaching time for the primary stage is determined empirically. However, tests have indicated that the primary stage leaching should preferable be terminated when the first tendency to hindered copper dissolution appears. After the primary stage for copper leaching, a wash/rinse is performed by addition of wash/rinse water. The rinse water may typically be acidic with low iron content, for example having a pH of about 2.5-5. The wash solution is diverted to a wash solution pond and thereafter subjected to an iron/aluminum/gypsum removal process.

The washed solids, i.e. the metallurgical residue, are thereafter subjected to a stage for selective removal of elemental sulfur. This stage is performed at a pH of 3.0 or higher, preferably a pH of at least 3.5. Selective removal of elemental sulfur may for example be made by biooxidation using a suitable biomass.

However, in accordance with a preferred embodiment, the stage for selective removal of elemental sulfur is an anaerobic sulfite leaching step. Sulfite should preferably be added in an amount which is over stoichiometric relative to the elemental sulfur demand, and the stage should preferably be made at a

temperature of at least 25 °C.

The leachate is diverted to a leachate pond and the bleed is subjected to a process for gypsum/aluminum removal. The solution may be recycled back to the leachate pond as shown in the figure.

After the stage for selective removal of elemental sulfur, the solids are subjected to a third stage for copper leaching at a lower pH by the action of ferric. It will be readily apparent to the skilled person that the stage for selective removal of elemental sulfur may be repeated after the third stage for copper leaching if necessary, and thereafter followed by yet another stage for copper leaching.

The heap leaching process disclosed in Figure 4 differs from previously known heap leaching processes in that it comprises a stage for selective removal of elemental sulfur, which is conducted at a pH of 3.0 or higher.

Since the elemental sulfur has been selectively removed, the leaching rate during the third stage for copper leaching will be substantially higher compared to if the elemental sulfur had not been selectively removed. In a conventional heap leaching process, layers of elemental sulfur and jarosite are formed on the surface of the chalcopyrite solids and thereby hinder the dissolution of the mineral.

However, by selectively removing elemental sulfur, formation of these types of surface layers during leaching is significantly reduced and the leaching rate thus improved. Even though Figure 4 relates to heap leaching of a chalcopyrite ore or concentrate, other types of sulfidic materials may also be heap leached with the process according to the present invention.

Moreover, the process according to the invention may also be a tank leaching process.

Example 1 - Selective removal of sulfur by bio-oxidation

Batch-wise tests for selective removal of sulfur by biooxidation have been carried out on bioresidues obtained from mesophilic bioleaching of a refractory gold concentrate produced from the from Petiknas North deposit.

The biooxidation of the mesophilic bio-residue was conducted batch-wise at about 10-20 % solids. A 1 -stage continuous culture (Dilution rate = 0.042 h^"1) grown on 1 g colloidal sulfur per liter at 47 °C, pH about 3.5 and about 4.5, respectively, were used as inoculum sources. Before the bio-residues were added, the continuous culture was switched to batch mode in order to oxidize the residual colloidal sulfur completely to sulfate.

The first batch was post-oxidized for 24 hours at about pH 3.5 and about 10 % solids. However, the actual found pulp solid content was about 17 %, since there had been an accumulation of gypsum in reactor. The final "true" residue was measured to about 0.17 % S°, corrected for the gypsum dilution given a sulfur removal of about 80 %. The complete sulfur removal profile was not determined.

The second batch was biooxidized for about 54 hours at about pH 4.5 and about 20 % solids. The obtained sulfur removal profiles are shown in Figure 5. As can been seen from the figure, the ultimate degree of sulfur conversion was about 80 % after 54 hours, giving a residual sulfur content of about 0.25 %.

A test where biooxidation was combined with subsequent sulfite leaching of the residual sulfur was also performed on a third batch. The third batch was first biooxidized for about 24 hours at a pH of about 4.5 and about 20 % solids. The pulp was subsequently transferred to a glass vessel for sulfite leaching at about 60 °C. The addition rate of Na2S2O₅ corresponded to about 2.5 times the

stoichiometric demand for complete dissolution by thiosulfate formation. The obtained sulfur removal profiles are shown in Figure 6. Example 2 - Selective removal of sulfur by sulfite leaching

Tests were performed on a bio-residue from Petiknas North refractory gold concentrate. The elemental sulfur content in the residue was about 0.5-0.6 %. Anhydrous sodium sulfite was used as sulfite reagent and NaOH for pH control.

The residue was leached at ambient temperature and about 46-47 °C at a pulp density of about 1 %. The sulfite was added in large excess versus the stoichiometric demand. The tests show that the contained bio-residue sulfur content was highly amenable to the leaching conditions and the dissolution rate is temperature dependent. At about 46 °C, pH 9.3 and 22 hours residence time, the elemental sulfur extraction was almost complete, approximately 99.7 %. The corresponding extraction at ambient temperature was approximately 86%. At about 47 °C, pH 7.8 and 1 .5 hours residence time, the extraction was

approximately 91 %.

Additional tests were undertaken to investigate leaching rate with respect to temperature and sulfite addition rate. All tests were carried out at a pulp solid content of approximately 18 % and slaked lime for pH control. The bio-residue was received from continuous bio-oxidation on Petiknas North low-grade refractory gold concentrate. The bio-oxidation was carried out in one stage at a residence time of 45 hours and at mesophilic temperature. The elemental sulfur content in the bio-residue was 0.59 %, based on 1 .5 hours extraction to a 20 mM

NaCN/acetone solvent followed by spectrometric determination via formation of ferric thiocyanate complexes.

The results are shown in Figures 7-10. It is clear that the sulfur extraction rate strongly depends on temperature and total sulfite addition.

Figure 7 illustrates the residual sulfur content in the bioresidue versus the leach time in hours. Leaching was performed at a temperature of about 45 °C and a pH of about 8.5. Sulfite was added in about 1 .1 times, 2.3 times and 3.5 times the stoichiometric demand (abbreviated in the figure to S.D.). Figure 8 illustrates the sulfur conversion versus the leach time in hours for the same conditions as in Figure 7.

Figure 9 illustrates the residual sulfur content in the bioresidue versus the leach time in hours. Leaching was performed at a temperature of about 65 °C and a pH of about 7.5. Sulfite was added in about 1 .1 times and 3.3 times the stoichiometric demand (abbreviated in the figure to S.D.). Figure 10 illustrates the sulfur conversion versus the leach time in hours for the same conditions as in Figure 9.

It was found that the formation of thiosulfate seemed to correspond stoichiometrically to the removal of elemental sulfur from the solids. However, at high reagent dosages there seemed to be a depletion of the sulfite ions from the solution by precipitation. The product may be CaSO3 or metal-SO3 compounds. The sulfur conversion profiles show however that such sulfite compounds are available for sulfur dissolution, but the availability is strongly affected by the temperature. This is rather clear form the tests at 65 °C where the degree of conversion at the lower dosage is approaching the one at the higher dosage by time.

Example 3 - Cyanide leaching test

Cyanide leaching tests were carried out on a bioresidue obtained from continuous 1 -stage bio-oxidation at about 37 °C on Petiknas low-grade refractory concentrate at 60 hours retention and about 15 % solids. The sulfur removal profile was not determined.

The bio-residue was filtered, thoroughly washed and re-pulped with water.

The re-pulped bio-residue was then split by a rotary divider into sub-samples, containing about 220-230 g of solids each on dry basis, which were used in the down-stream tests. The head analysis of the bioresidue is given in Table 1 . The elemental sulfur analysis was made by gravity determination after extraction to carbon disulfide.

Three tests, Test nos. A-C, were performed as briefly described in Table 2 and the calculated head analyses of Au, Ag and S° is summarized in the same table. The head and residue analysis were made by extraction to a 20 mM

NaCN/acetone solvent followed by spectrometric determination via formation of ferric thiocyanate complexes. The calculated elemental sulfur head assays are based on residual elemental sulphur after cyanidation and the formed thiocyanate.

It was found that in Tests nos. A and C, i.e. when a major part of the elemental sulfur had been removed from the bioresidue, the initial leaching rate was significantly faster than the initial leaching rate in Test no. B, where the bioresidue was subjected directly to cyanidation without prior removal of the cyanide reactive sulfur content. In fact, it was found that the ultimate gold recoveries are already achieved after about 5-10 hours for Test nos. A and C, while in Test no. B more than 24 hours was needed. This can be seen in Figure 1 1 , which shows the gold in leach tails versus the leach time for Test nos. A-C. Table 1

Table 2.

In Figure 12a, the obtained NaCN consumption profile is shown. It is clear from the figure that the sodium cyanide consumption is significantly less for Test nos. A and C, compared to Test no. B. Moreover, it should be noted that the actual cyanide consumption will depend on the leaching time, since free cyanide has to be maintained to control the leaching rate of gold and silver. To obtain the ultimate gold recovery in Test no. B, 24 hours leaching was needed in batch leaching, as mentioned above, and the cyanide consumption for 24 hours was about 12 kg NaCN per ton bioresidue as shown in Figure 12a. However, to achieve the ultimate gold recoveries in Test nos. A and C, only about 1 -2 kg NaCN per ton was consumed since the required leaching was only about 5-10 hours. Figure 12b show the thiocyanate profile for the three tests. It is clear that the thiocyanate formation was extensive in Test no. B. The calculated head assays indicate that SCN^" is mainly formed from the elemental sulfur content.

Additional tests were performed on the mesophilic bioresidue mentioned above for three different processes, Test nos. D-F, according to the invention, and compared to a reference sample, Test no. G, wherein elemental sulfur had not been selectively removed, with the object to determine if there is an agreement between the elemental sulfur content of the bioresidue and the thiocyanate formation during cyanide leaching. The tests nos. D-F are summarized below in Table 3 as well as the elemental sulfur content in the calculated head grade, the assayed head grade and the final residue. The calculated heads are based on the sulfur found as thiocyanate and the sulfur in the final cyanidation residues.

Table 3.

As previously mentioned, there has been a debate through the years regarding the role of the enzyme rhodanese and how it affects the cyanide consumption. One theory has been that the rhodanese activity from the biomass contributes significantly to the loss of cyanide in the gold extraction plant by enhancing the thiocyanate formation by cleaving S-S bonds. Another theory has however been that it is unlikely that the rhodanese activity is responsible for the excessive cyanide wastage at the high pH values associated with the cyanidation process. The very good correlation between thiocyanate formation and elemental sulfur removal found in the above described tests, and shown in Table 3, support the theory that thiocyanate formation is purely a result of chemical reaction between the S° and cyanide ions.

Example 4 - Selective removal of sulfur by sulfite leaching of copper concentrate

A batch-wise bioleaching test was performed on a dirty copper-lead concentrate from Maurliden. The head analyses are shown in Table 4.

Table 4.

The test was carried out with a bacterial culture, which had grown on a pyrite concentrate from Aitik at about 37 °C and pH about 1 .5, in batch mode. The Cu-Pb concentrate was added stepwise, into about 3 liters of an active leach solution from the pyrite concentrate. The initial iron concentration was about 5 g/l and dominated by ferric iron. The final pulp density was about 10 % of solids, by weight. The pH was stabilized at about 1 .5, without need for pH control. The obtained leaching profile is shown in Figure 13. The shown copper recovery profile in Figure 13 is based on total added copper and solution analyses.

The leaching rate was initially slow. At day 12, the batch leaching was stopped and the solution was separated from the solids. The solution was returned to the bioreactor to be aerated in order to maintain the activity. The redox potential was increasing gradually to >600 mV, which indicated that the culture was still active. The washed solids were subjected to sulfite leaching at 65 °C for 24 hours. Sodium sulfite was added in excess, for the dissolution of S° as thiosulfate. The initial pH was about 9.5 using NaOH addition. By sulfite leaching, the cyanide reactive sulfur content was reduced from about 1 .36 % to about 0.10 %.

After sulfite leaching, the residue was returned to the bioreactor, where it was re-pulped with the recycled bioleachate. As can be seen in Figure 13, the redox potential dropped drastically and was never restored. The observed copper extraction rate was very slow. The cause of the inactivity was probably due to the culture being poisoned by the dissolution of sulfites, which were present in the residue, after sulfite leaching. After 5 days, 100 ml of fresh bioleachate from a pyrite bioleaching reactor was added and the copper extraction rate suddenly increased drastically after one day. The leaching was terminated after one further day, to enable assaying of final residue. The copper analysis of final solution indicated about 47 % recovery to solution based on the head assay. The assay of the final residue is given in Table 5. In Table 6, the calculated recoveries of metals to solution, based on the assumption of negligible dissolution of lead, are given.

The concentration of cyanide reactive sulfur in the final bioresidue was about 0.7 %, which was about half the concentration found for the bioresidue from the first bioleaching stage.

Table 5.

The test was repeated in order to confirm the positive effect of the elemental removal on the copper extraction rate. The obtained leaching profile shown in Figure 14. The shown copper recovery profile is based on the actual concentrate added to the reactor. Analyses of the solution show a very rapid extraction rate for the first additions of concentrate. At 48 % of the total addition, about 85 % of the copper recovery to solution was achieved within 5 days. When the remainder additions of concentrate were added the extraction rate decreased significantly.

After 7 days, the bioleaching was stopped and the solution was separated from the solids and leached by sulfite at about 65 °C. The bioleach solution was returned to the bioreactor and the redox potential was restored to >700 mV.

As in the first test, the sulfur content in the residue was about 1 .4 % when the bio-oxidation was stopped. The sulfite leaching reduced the S° content also in this test to about 0.1 %. After re-pulping the residue in the restored bio-leachate, the copper dissolution increased rapidly, as in the first test. After totally 25 days of bioleaching, the copper recovery was about 90 %.

The bio-residue was regularly sampled for S° analysis during the course of bioleaching, after sulfur removal. It is interesting to note that S° content in the bioresidue seems to stop at about 0.6-0.7 %.

Claims

1 . Process for recovering a valuable metal from a sulfidic material comprising said valuable metal, the process comprising the steps of:

- subjecting the sulfidic material to at least a first hydrometallurgical step and subjecting the discharged slurry to a solid/liquid separation step thus obtaining a metallurgical residue,

- subjecting the metallurgical residue to a step for selective removal of

elemental sulfur, wherein said step for selective removal of elemental sulfur is conducted at a pH of 3.0 or higher, and

- subjecting the metallurgical residue to a final metal recovery step.

2. Process according to claim 1 , wherein said first hydrometallurgical step is a bioleaching step.

3. Process according to any of claims 1 or 2, wherein said step for selective removal of elemental sulfur is a biooxidation step.

4. Process according to claim 3, wherein the biooxidation step for selective removal of elemental sulfur is performed at a pH of 3.5-6, preferably at a pH of

3.8-5.5.

5. Process according to any of claims 1 or 2, wherein said step for selective removal of elemental sulfur is a sulfite leaching step.

6. Process according to claim 5 wherein the sulfite leaching is performed at a pH of 3.5 or higher, preferably at a pH of 3.5-8.

7. Process according to any of claims 5 and 6 wherein the sulfite leaching step is performed at a temperature of above 25 °C, preferably above 40 °C.

8. Process according to any of the claims 5 to 7 wherein sulfite is added to the sulfite leaching step in an amount which is over-stoichiometric relative to the elemental sulfur demand.

9. Process according to claim 8 wherein sulfite is added in an amount of at least twice the stoichiometric demand of elemental sulfur.

10. Process according to any of claims 5 to 9 wherein the metallurgical residue is subjected to a biooxidation step at a pH of at least 3 after said sulfite leaching step.

1 1 . Process according to any of claims 5 to 9 wherein the metallurgical residue is subjected to a thiosulfate leaching step after said sulfite leaching step.

12. Process according to any of the preceding claims wherein the sulfidic material is an ore or a concentrate comprising arsenopyrite and/or iron pyrite.

13. Process according to any of claims 1 to 1 1 wherein the sulfidic material is chalcopyrite.