FI125027B - Menetelmä metallien talteenottamiseksi niitä sisältävästä materiaalista - Google Patents

Description

Method for recovering metals from material containing them FIELD OF THE INVENTION

The invention relates to a method for recovering metals from material containing them.

BACKGROUND OF THE INVENTION

The conventional pyrometallurgical way to process nickel sulphide concentrate is to process the concentrate in a smelter into nickel matte and further in a converter into high-grade nickel matte. The oxidation degree of the concentrate feed i.e. the amount of oxygen to be fed in determines the iron content of the matte generated on the furnace floor as well as the nickel and copper content of the slag. The optimal ratio between the iron content of the matte and the nickel and copper content of the slag can be controlled by adjusting the ratio of the concentrate and oxygen fed into the furnace (Nm3 O2/1 cone.).

Nickel sulphide concentrate and thus also high-grade nickel matte usually always also contains copper and so nickel recovery from matte is also the separation of nickel from copper. The hydrometallurgical processing of nickel-copper matte is described in several patent publications such as for instance US 4,323,541 and US 5,628,817.

When sulphidic bulk concentrate is concerned, in which in addition to nickel the amount of copper and possibly that of cobalt is considerable, it has to be considered carefully how to recover the different metals in the simplest and most economical way possible.

Earlier proposals have included the slow cooling of high-grade nickel matte, so that nickel and copper separate into their own fractions in the cooled matte. Slow cooling is also a question of cost and possibly selectivity also suffers. US patent publication 5,628,817 describes a method for recovering nickel and copper from high-grade nickel matte in a sulphate-based process, in which there is first two-stage atmospheric leaching followed by two-stage pressure leaching. The final recovery of copper and nickel takes place in the corresponding electrowinning. In the first atmospheric leaching stage, Ni-Cu matte is leached by means of oxygen and copper sulphate solution, so that the nickel dissolves and the soluble copper is precipitated. The precipitate from the first leaching stage is routed to the second atmospheric leaching, where leaching occurs in oxidizing conditions with the anolyte containing sulphuric acid from nickel electrowinning. Besides nickel, copper also dissolves and the solution is fed back to the first leaching stage. The precipitate from the second atmospheric leaching is routed to the first pressure leaching stage, where the remaining nickel is leached by means of copper sulphate. The solution that is generated is routed via iron removal to the second atmospheric leaching and the precipitate to the second pressure leaching. The precipitate now contains mostly secondary sulphides of copper, which are leached with the anolyte from copper electrowinning. The precipitate of the second pressure leaching contains any precious metals there may be. The solution exiting the second pressure leaching contains copper sulphate and impurities, which are removed before copper electrowinning. A method is disclosed in US patent publication 6,039,790, in which nickel is recovered from two different nickel mattes in the context of the same process, mattes from smelting furnace and matte from an electric furnace. The iron content of the electric furnace matte is far higher than that of the smelting furnace matte, which is high-grade nickel matte. The leaching of the fine-grade nickel-copper smelting furnace matte is carried out in two atmospheric and one pressure leaching stage and that of the electric furnace matte in one atmospheric leaching stage with a sulphate-based solution. The smelting furnace matte is leached in the first atmospheric leaching stage with the nickel sulphate solution containing copper sulphate exiting the second atmospheric leaching stage. The nickel sulphate solution that is generated is routed to nickel electrowinning after solution purification (cobalt removal). The precipitate of the first leaching stage is routed to the second leaching stage, in which leaching takes place with the nickel sulphate solution exiting the leaching of the electric furnace matte and with anolyte from nickel electrowinning. The precipitate of the second leaching stage is routed to a pressure leaching stage, where leaching occurs with nickel electrowinning anolyte. In the pressure leaching stage, the secondary nickel sulphide generated in earlier stages dissolves and copper is precipitated. The copper precipitate, which also contains precious metals, is routed for example to pyrometallurgical treatment. The nickel sulphate solution formed in pressure leaching, which also contains ferrous sulphate, is routed to the leaching of the electric furnace matte. Some neutralizing agent is also fed into the leaching stage in order to precipitate the iron as jarosite.

The leaching of nickel-copper matte in a chloride environment is described for example in US patent publications 3,880,653 and 3,975,189. In these methods the matte is leached and nickel recovered as metal in chloride electrowinning. Copper is precipitated from solution before electrowinning and the precious metals contained in the matte (gold and PGM) do not dissolve but remain in the anode slime from electrowinning and can be recovered from there. US patent publication 7,736,606 describes the leaching of nickel concentrate or matte as a chloride-based leach. Leaching is performed in conditions where the soluble sulphur from the concentrate forms hydrogen sulphide, which is removed from the solution. Magnesium chloride and some strong oxidant such as chlorine or hypochlorite are also routed to leaching in order to improve the nickel leaching yield. Other valuable metals in the concentrate or other raw material used, such as copper and cobalt, as well as iron, also dissolve in the leaching conditions. Gold and PGM are also partially dissolved. The dissolved gold and PGM are recovered from the solution containing valuable metals and the nickel chloride solution is subsequently subjected to solution purification. Solution purification takes place by means of extraction and precipitation, whereby first copper is extracted from the solution, then iron is precipitated and finally nickel/cobalt extraction is performed. Solutions rich with regard to each metal are routed to recovery of the corresponding metal. A method is described in US patent publication 6,428,604 for the recovery of nickel and cobalt from a sulphidic flotation concentrate. The first leaching stage is carried out at atmospheric pressure with a sulphate-chloride solution, into which chlorine is also fed. The chloride concentration of the solution is 2 - 40 g/l. The slurry (solution + solids) of the first leaching stage is routed to pressure leaching, where the oxidant used is oxygen. The presence of chloride in leaching promotes leaching and prevents the formation of sulphate instead of elemental sulphur. After solids-liquid separation the solution is subjected to copper removal either by precipitation or by extraction. Next, iron is removed from the solution by precipitation using lime, and impurities such as zinc, lead and the remaining copper are removed from the solution in solution purification. The purified solution containing nickel and cobalt is subjected to nickel-cobalt extraction to separate the cobalt from the solution. The chloride-containing nickel sulphate solution is routed to nickel recovery through elec trowinning. Metallic nickel and chlorine gas are formed in electrowinning, and the chlorine is routed to the first leaching stage.

PURPOSE OF THE INVENTION

The advantage of the method accordant with the invention over conventional methods is superior manageability of sulphur and in particular of iron.

SUMMARY OF THE INVENTION

The invention relates to a method, whereby valuable metals, as well as any precious metals there may be, are recovered from a material containing them, such as the mixed matte formed in a smelter. The valuable metals of material formed in a smelter are leached with an acidic solution containing sulphate and chloride, from which each metal is separated by means of solvent extraction. In leaching any precious metals contained in the material remain undissolved in the leach residue, from which they can be leached with a solution containing hydrochloric acid in oxidizing conditions and separated from the solution by solvent extraction.

The invention relates to a method for leaching the fine-grained mixed matte formed in a smelter that contains iron and valuable metals, and separating the valuable metals from the leaching solution. The leaching stage of the valuable metals in the matte is carried out at atmospheric pressure and in oxidizing conditions with an acidic leaching solution containing sulphate and chloride and the valuable metals are separated from the leaching solution by solvent extraction.

According to one embodiment of the invention, the valuable metals in the mixed matte are nickel and copper. According to a second embodiment of the invention, in addition to nickel and copper, the valuable metal is at least one of the following: cobalt and silver.

In one embodiment of the invention, the iron in the mixed matte is leached and precipitated during the leaching stage, after the leaching stage solids-liquid separation is performed, whereby the iron-containing leach residue is separated from the solution containing the valuable metal.

According to one embodiment of the invention, separation of the valuable metals from the leaching solution is carried out by solvent extraction in the following order: silver, copper, cobalt and nickel.

In the method accordant with the invention, the amount of chloride in the leaching solution is typically 150 - 200 g/l and the amount of sulphate 70 -100 g/l.

One embodiment of the invention is that, in addition to valuable metals, there are precious metals in the mixed matte, the precious metal being at least one of the following: gold, palladium and platinum; the precious metals remain undissolved in the mixed matte leaching stage and remain in the iron-containing leach residue.

In one embodiment of the invention, the mixed matte leach residue is routed to a second leaching stage, where first iron is leached from the leach residue in sulphate- and chloride-based leaching, after which the slurry that is formed is subjected to solids/liquid separation in order to separate the iron-containing solution and the precious metal precipitate from each other. The iron-containing solution is neutralised to a value of 2 - 2.5 to precipitate the iron from the solution.

According to one embodiment of the invention, leaching of the iron-depleted precious metal precipitate is carried out with concentrated hydrochloric acid solution in oxidizing conditions at atmospheric pressure. The hydrochloric acid solution is over 5 molar, preferably 6-8 molar.

According to one embodiment of the invention, the precious metals dissolved in the hydrochloric acid solution are separated from the precious metal product solution by solvent extraction in the order gold, palladium, platinum.

According to one embodiment of the invention, at least part of the precious metal-depleted hydrochloric acid solution is fed back to the precious metal leaching stage.

According to one embodiment of the invention, part of the precious metal-depleted hydrochloric acid solution is fed to the valuable metal leaching stage.

LIST OF DRAWINGS

Figure 1 presents one embodiment of the invention diagrammatical- iy-

DETAILED DESCRIPTION OF THE INVENTION

Sulphidic concentrate containing nickel and copper as well as other valuable metals is subjected first to smelting treatment for example in a flash smelting furnace or other equivalent smelter to form NiCu matte. Smelter matte may also be formed in an Ausmelt furnace or an electric furnace or may be a combination of the two. One method to manufacture high-grade nickel smelter matte is described in CA patent publication 2,008,167, according to which matte and slag are formed in a suspension smelting furnace, the slag is routed on to electric furnace treatment and the matte formed there is fed into the suspension smelting furnace, so that treatment in two furnaces results in a single matte. NiCu matte may also be the product of a scrap smelter, Kaldo furnace or TROF furnace. When the starting material of the smelter is a sulphidic concentrate, the iron it contains remains mainly in the slag formed in the furnace, whereas the nickel, copper and other valuable metals, such as silver and cobalt plus precious metals, such as gold, platinum and palladium, are concentrated in the matte formed in the furnace. Since the matte contains several metals, it is known as mixed matte. The amount of valuable metals in the matte may vary. It is typical of the leaching method accordant with the invention, but not essential, that in addition to valuable metals the matte also contains precious metals.

The method accordant with the invention is illustrated by flow chart 1. The mixed matte formed in the smelter is granulated, milled and elutriated into a raffinate, which is recycled to the first leaching stage from the tail end of the process. The raffinate is a chloride-sulphate solution containing sulphuric acid and depleted of valuable metals, in which the chloride concentration of the solution in particular has been adjusted to close to saturation and the sulphate concentration is also high. The chloride and sulphate concentration of the raffinate is controlled by feeding a hydrochloric acid-containing solution and sulphuric acid into the leaching stage as required. Typically the chloride concentration is about 150 - 200 g/l and the sulphate concentration about 70-100 g/l. The leaching stage consists of several agitated reactors. Leaching is performed at atmospheric pressure and elevated temperature (85 - 103°C) in oxidizing conditions. The oxidizing conditions are achieved by feeding an oxygen-containing gas into at least some of the leaching reactors. Leaching is based mostly on the oxidizing capability of divalent copper, whereby copper is reduced to monovalent, as well as on the concurrent oxidation/reduction reaction of iron.

During the first leaching stage, the valuable metals in the mixed matte are leached out, i.e. nickel, copper, cobalt as well as any silver that may be in the matte. At the same time, part of any lead and zinc that may be in the mixed matte may also dissolve. Calcium and magnesium components also dissolve, though there are very small amounts of them, because calcium and magnesium compounds primarily go into the slag of the smelting process. During leaching, iron dissolves and is precipitated mostly as hematite and goe-thite, as the solution is neutralized at the end of the leach to a pH value of 2 -2.5. If the mixed matte contains arsenic, this is precipitated along with the iron as ferric arsenate. The sulphur of the sulphidic mixed matte is oxidized partially to sulphate and part remains in the elemental form. Of the metals it is mainly the precious metals that remain undissolved, such as gold, platinum and palladium, if they are present in the mixed matte. When the amount of precious metals is significant, they are leached in a second leaching stage.

The slurry formed in the first leaching stage is routed to solids/liquid separation, which takes place for instance in a thickener (not shown in detail in the drawing). When there is a significant amount of precious metals, the underflow of the thickener is filtered and routed to the second leaching stage to leach the precious metals. The thickener overflow forms the valuable metal-containing product solution, which is routed to the solution purification stage.

The solution purification stage consists of several solvent extraction stages, into which the valuable metal-containing product solution is routed. When the valuable metals are nickel, copper, cobalt and silver, solution purification is performed as four consecutive extraction stages. Each extraction stage includes the actual extraction, scrubbing and stripping of the extraction solution, which are the conventional stages of extraction. If the final recovery of the valuable metal occurs by electrowinning, the aqueous stripping solution used is a solution of sulphuric acid. In such a case, the valuable metal in question can be routed to electrowinning in sulphate form, which makes electrowinning simple. For the sake of simplicity, the final treatment for each metal is marked electrowinning (EW) on the diagram, although of course it may also be precipitation.

When the product solution includes silver, it is removed from the solution first. For example, a solvating triisobutylphosphine sulphide extractant, for instance CYANEX 471X, which has been modified with D2EHPA extractant (di-2-ethylhexyl phosphoric acid), may be used as the organic extraction solution. In the extraction stage, the silver of the product solution is transferred to the organic extraction solution and the silver-depleted product solution is rout ed to the next stage of solution purification. The stripping of silver from the organic solution preferably takes place with a stabilized aqueous solution containing sodium thiosulphate. Silver is recovered from the stripping solution by a method suitable for the purpose, either by reducing precipitation or electrolyti-cally.

Next, the product solution is subjected to copper removal. If the amount of silver in the product solution is so low that it does not need to be removed separately, copper removal is performed first. It is preferable to extract copper for instance with a hydroxyoxime reagent, such as LIX84 in a pH range below 3, whereby nickel is not extracted along with it. After copper extraction the copper-poor product solution is routed to the next stage of solution purification. The post-extraction copper electrowinning raffinate, i.e. sulphuric acid solution is used as the aqueous solution for copper stripping, and the copper-rich sulphuric acid solution is routed to electrowinning to produce metallic copper.

The next purification stage for the product solution is cobalt removal, which also takes place by means of extraction. Since the solution used for matte leaching contains an abundance of chloride in addition to sulphate, there are significant amounts of cobalt in the solution as a tetrachloro complex. It is preferable to extract the CoCU2' anion using for example a tertiary amine as organic extractant, of which one brand is Alamine 336. Extraction is carried out at a pH value of about 3. The stripping solution is preferably sulphate-based and cobalt is recovered from it by an appropriate method either chemically by precipitation or by electrowinning.

The product solution containing chloride and sulphate, which has been subjected to solution purification stages in which the other valuable metals have been removed from the solution, is routed to nickel extraction. As stated above, in addition to valuable metals the solution contains calcium, magnesium and other impurities and therefore it is most advantageous to separate nickel from the product solution by means of a separate extraction stage. The pH of the product solution is raised to a value of 3.5 - 4 and a hydroxyoxime reagent such as LIX84 is used as organic extraction solution, in which another extraction reagent is also present such as branched C-10 tertiary carboxylic acid, of which one brand is Versatic 10. The latter acts as a synergistic extractant and brings about the extraction of nickel in an environment that is two pH units more acidic than by using a hydroxyoxime reagent alone, so that it saves considerably on solution neutralization costs. The aqueous solution of nickel stripping that is used is a sulphuric acid solution, which is preferably the anolyte from nickel electrowinning and which solution is routed to nickel electrowinning to produce metallic nickel.

All the valuable metals have been removed as described above from the product solution of mixed matte leaching, so it can be termed the raffinate. Of course, a sidestream of the raffinate can be taken in order to remove dissolved impurities. The raffinate is still acidic, but its chloride concentration has been reduced and the chloride concentration has to be raised before the raffinate is fed back to the leaching stage. At the same time the acid concentration of the raffinate has to be adjusted to the desired level. If the leaching process is carried out at the same facility as the smelting process, the sulphur dioxide formed in concentrate smelting is generally processed into sulphuric acid and therefore the sulphuric acid formed in this way can be used in the control of the acid concentration.

If the precious metals content of the mixed matte formed in smelting is so high that a separate leaching process for them is financially viable, it is advantageous to perform leaching as a second leaching step, which is chloride-based with regard to precious metals leaching. Since the amount of precious metals is small, the separate chloride circuit is also made small. The precious metal-depleted hydrochloric acid exiting the precious metals leaching circuit can not only be recycled back to precious metals leaching, but also used for the control of the chloride concentration of the first leaching stage mentioned above. The leach residue from the mixed matte leaching stage is mostly iron precipitate, which also contains sulphur and the precious metals that have remained undissolved. The amount of sulphur in the leach residue is small and it can be removed by known methods. The thickened and filtered leach residue from the first leach is routed to the precious metals leaching and recovery stage. When the leach residue is a filter cake of precipitate, it minimizes the transfer of other metals to this stage.

In the second leaching stage, the treatment of the valuable metal-containing leach residue begins with iron removal. The leach residue is elutriated into an aqueous solution containing chloride and sulphuric acid. The iron in the residue dissolves easily as sulphate, but the precious metals do not dissolve in this stage. The slurry is subjected to solids/liquid separation, in which the precious metals remain in the precipitate. The iron-containing solution is routed to the iron precipitation stage, which is performed in the usual way for example by means of a lime compound. Neutralization is carried out at a pH value of about 2 - 2.5, so that the iron and any arsenic that may be in the solution are precipitated from the solution. If necessary, the iron precipitate that is formed can be returned to the smelter, especially if its arsenic content is low. In smelting treatment the iron is converted to slag.

After iron removal, the precipitate containing precious metals is routed to the actual precious metals leaching stage. In the leaching stage the precipitate is leached with concentrated hydrochloric acid of over 5 molar, preferably 6-8 molar, in oxidizing conditions at atmospheric pressure and temperature. Some strong oxidant such as hydrogen peroxide or the equivalent is used as oxidant. Hydrogen peroxide reacts with hydrochloric acid and forms reactive chlorine gas for use locally in the process. Obviously, closed reactors are used for leaching, which means that the reactor is equipped with a cover but however works at atmospheric pressure. Gold, platinum and palladium dissolve in the leaching conditions.

The precious metals are recovered from the precious metal-containing product solution. This takes place preferably by separating each metal individually from the solution by means of solvent extraction. When there is gold in the solution, gold extraction is carried out first. The preferred organic extractant is 2,2,4-trialkyl-1,3-pentanediol diester and even more advantageous is a mixture of the above ester and a branched long-chain alcohol. The CAS number of one suitable diester is 6846-50-0. When using the extraction reagent mentioned above, plain water can be used as the aqueous solution in stripping. Gold can be reduced to metal from the aqueous solution by means of oxalic acid for instance.

After gold, palladium is extracted from the solution containing precious metals. A solvating tri-isobutylphosphine sulphide such as CYANEX 471X for example can be used as organic extraction reagent, from which palladium stripping can be performed using stabilized sodium sulphate. Palladium can be precipitated from the stripping solution for instance by reducing it with sodium borohydride NaBH4.

After palladium extraction, the only precious metal remaining in the precious metals-containing solution is platinum, and a suitable extractant for platinum extraction is a secondary amine, such as for example the amine known by the product name Amberlite LA-2. Platinum stripping can be done for instance with an aqueous alkaline solution as presented in patent publication US 4,041,126. The precipitation of platinum from the stripping solution can be done for example with ammonium chloride, whereupon the product obtained is ammonium chloroplatinate, as described in the above-mentioned patent publication. As an alternative extractant, a solvating reagent may be used, for example the reagent with the product name Cyanex 923, from which platinum can be stripped with an aqueous solution of about 30% nitric acid, for instance. Both the precipitated palladium and the platinum are recovered as extremely pure metals and are commercial goods as they are.

The hydrochloric acid solution depleted of precious metals may contain small amounts of copper, nickel and cobalt. The relevant metals are precipitated from the solution and returned to the mixed matte leaching circuit (not shown in detail in the diagram). At least some of the hydrochloric acid solution depleted of precious metals is recycled back to the leaching of precious metal precipitate. As stated above, if necessary some of the solution is also routed to the leaching of mixed matte.

Obviously, the details of the method described above may be varied, for instance, other extractants than those mentioned above may be used. The invention and its embodiments are thus not restricted to the example models described above, and may vary within the framework of the patent claims.

Claims (14)

1. Menetelmä sulatusuunissa muodostetun ja hienojakoiseksi jaetun rautaa ja arvometalleja sisältävän sekakiven liuottamiseksi ja arvometallien erottamiseksi liuotusliuoksesta, jolloin kiven arvometallien liuotusvaihe suoritetaan atmosfäärisessä paineessa ja hapettavissa olosuhteissa sulfaattia ja kloridia sisältävällä happamalla liuotusliuoksella ja arvometallit erotetaan liuotuksessa muodostetusta tuoteliuoksesta neste-nesteuutolla ja jolloin sekakiven arvometalleja ovat nikkeli ja kupari.

2. Patenttivaatimuksen 1 mukainen menetelmä, jolloin arvometalleja on nikkelin ja kuparin lisäksi ainakin yksi joukosta koboltti ja hopea.

3. Jonkin edellisen patenttivaatimuksen menetelmä, jolloin sekakiven rauta liuotetaan ja saostetaan liuotusvaiheen aikana, liuotusvaiheen jälkeen suoritetaan kiintoaine-neste-erotus, jolloin rautapitoinen liuotusjäännös erotetaan arvometallipitoisesta tuoteliuoksesta.

4. Jonkin edellisen patenttivaatimuksen menetelmä, jolloin arvometallien erotus liuoksesta suoritetaan neste-nesteuutolla järjestyksessä hopea, kupari, koboltti ja nikkeli.

5. Jonkin edellisen patenttivaatimuksen menetelmä, jolloin kloridin määrä liuotusliuoksessa on 150 - 200 g/l.

6. Jonkin edellisen patenttivaatimuksen menetelmä, jolloin sulfaatin määrä liuotusliuoksessa on 70- 100 g/l.

7. Jonkin edellisen patenttivaatimuksen menetelmä, jolloin sekaki-vessä on arvometallien lisäksi jalometalleja; jalometallin ollessa ainakin yksi joukosta kulta, palladium ja platina; jalometallit jäävät liukenematta sekakiven liuotusvaiheessa ja jäävät rautapitoiseen liuotusjäännökseen.

8. Patenttivaatimuksen 7 mukainen menetelmä, jolloin sekakiven liuotusjäännös johdetaan toiseen liuotusvaiheeseen, jossa ensin liuotetaan rauta liuotusjäännöksestä sulfaatti- ja kloridipohjaisessa liuotuksessa, jonka jälkeen lietteelle suoritetaan kiintoaine/neste-erotus rautapitoisen liuoksen ja jalometallisakan erottamiseksi toisistaan.

9. Patenttivaatimuksen 8 mukainen menetelmä, jolloin rautapitoinen liuos neutraloidaan arvoon 2 -2,5 raudan saostamiseksi liuoksesta.

10. Patenttivaatimuksen 8 tai 9 mukainen menetelmä, jolloin rauta-köyhän jalometallisakan liuotus suoritetaan väkevällä suolahappoliuoksella hapettavissa olosuhteissa atmosfäärisessä paineessa.

11. Patenttivaatimuksen 10 mukainen menetelmä, jolloin suolahap-poliuos on yli 5 molaarinen, edullisesti 6-8 molaarinen.

12. Patenttivaatimuksen 10 tai 11 mukainen menetelmä, jolloin suo-lahappoliuokseen liuenneet jalometallit erotetaan jalometallien tuoteliuoksesta neste-nesteuutolla järjestyksessä kulta, palladium, platina.

13. Jonkin patenttivaatimuksen 8-12 mukainen menetelmä, jolloin ainakin osa jalometalliköyhästä suolahappoliuoksesta syötetään takaisin jalo-metallien liuotusvaiheeseen.

14. Jonkin patenttivaatimuksen 8-13 mukainen menetelmä, jolloin osa jalometalliköyhästä suolahappoliuoksesta syötetään arvometallien liuotusvaiheeseen.