EP2350327A1 - Method for treating nickel laterite ore - Google Patents

Method for treating nickel laterite ore

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Publication number
EP2350327A1
EP2350327A1 EP09828687A EP09828687A EP2350327A1 EP 2350327 A1 EP2350327 A1 EP 2350327A1 EP 09828687 A EP09828687 A EP 09828687A EP 09828687 A EP09828687 A EP 09828687A EP 2350327 A1 EP2350327 A1 EP 2350327A1
Authority
EP
European Patent Office
Prior art keywords
acid
stage
ore
laterite
routed
Prior art date
Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
Withdrawn
Application number
EP09828687A
Other languages
German (de)
French (fr)
Other versions
EP2350327A4 (en
Inventor
Mikko Ruonala
Bror Nyman
Jaakko Leppinen
Ville Miettinen
Teppo Riihimäki
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
Outotec Oyj
Original Assignee
Outotec Oyj
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Application filed by Outotec Oyj filed Critical Outotec Oyj
Publication of EP2350327A1 publication Critical patent/EP2350327A1/en
Publication of EP2350327A4 publication Critical patent/EP2350327A4/en
Withdrawn legal-status Critical Current

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Classifications

    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/005—Preliminary treatment of ores, e.g. by roasting or by the Krupp-Renn process
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
    • C22B23/0469—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods by chemical substitution, e.g. by cementation
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0476—Separation of nickel from cobalt
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

Definitions

  • the invention relates to a method for processing laterite ores so that the metals contained in the laterites are turned into water-soluble form for the recovery of valuable metals, such as nickel and cobalt.
  • metals such as nickel and cobalt.
  • Different types of nickel laterites are processed at the same time without being separated according to their iron and/or magnesium content.
  • laterites are pre- treated with concentrated mineral acid so that the metals contained in the laterites react to form water-soluble salts, the silicates contained in the laterites are partially decomposed and post-leaching liquid-solid separation becomes easier than it was previously.
  • the water vapour generated in the ore and acid reaction stage is utilised in ore drying and the unreacted mineral acid is recycled to the front end of the process.
  • a method is disclosed in US patent 4125588 in which nickel laterite ore is pre-treated with concentrated acid before nickel leaching.
  • laterite ore is dried so that the moisture content of the ore is less than 1 %.
  • the dried ore is ground to a particle size range of 65 - 100 mesh.
  • the ground ore is mixed into concentrated acid at a mass ratio of about 1 :1.
  • the metal sulphation reactions are initiated by adding water to the mixture containing laterite and acid in a ratio of 3 - 40% of the mass of the laterite.
  • the sulphated metals are leached into water.
  • US patent application 2006/0002835 describes a method in which laterite leaching takes place in two stages.
  • the laterite ore is mixed with concentrated sulphuric acid.
  • the ore/acid mixture is slurried with water and then leached, so that the nickel and cobalt dissolve.
  • the amount of sulphuric acid used in pre- treatment is stoichimetric with regard to the non-ferrous metals in the ore, but not the iron. According to example 5, the acid/ore ratio is 0.65. A little excess acid is advantageous so that a small amount of iron also dissolves, because this achieves the maximal dissolution of nickel and cobalt.
  • the leaching stage of the method is implemented either at a temperature of 95 - 105° C or in an autoclave at a temperature which is a maximum of 150° C and in which the pressure corresponds to the saturated steam pressure.
  • some suitable reductant is added to the leaching stage, such as sulphur dioxide.
  • sulphur dioxide is added to the leaching stage.
  • the effectiveness of the method requires that there is a significant proportion of easily-soluble saprolitic laterite in the laterite, so that it can function in the ferric iron precipitation range.
  • the purpose of the invention presented here is to eliminate the problems that have arisen in earlier methods and to achieve a method for leaching laterites in a way that in particular solids separation from the solution does not cause problems.
  • the invention also aims to improve the economics of the laterite process by utilising the heat generated in the reactions and recovering the acid that is not consumed in the reactions.
  • the invention relates to a method for processing nickel laterite ore to facilitate the recovery of nickel and cobalt and liquid-solid separation.
  • a) crushed laterite ore is dried using the steam from a later process stage, b) the dried ore is subjected to dry grinding, c) the fine ore is routed to an acid treatment mixing stage that takes place using concentrated mineral acid, in which the ore is mixed homogenously into the acid, the amount of which is at least stoichiometric with regard to the metals contained in the ore, d) the mixture of ore and acid is fed to the acid treatment reaction stage, which takes place at ambient pressure and a temperature of between 15O 0 C and the boiling point of the acid, whereby the acid and ore react with each other; the water vapour formed in the reaction is recovered and recycled for use in ore drying, e) the acid-treated laterite ore is routed to acid recovery, where the acid that remained unreacted in the reactions of the acid treatment mixing stage is recovered by evaporation,
  • the particle size of the crushed laterite ore is in the region of 90% below 10 mm.
  • the acid and laterite are mixed into a homogenous mixture at a temperature where the metals do not yet substantially react with the acid.
  • the metals of the laterite ore are made to form water-soluble salts of the mineral acid.
  • the mineral acid is at least one of the following: sulphuric acid, nitric acid, or hydrochloric acid, or a mixture of at least two of them.
  • the concentration of the mineral acid is preferably in the order of 70 - 98 wt.%.
  • the reaction stage is sulphation.
  • the mixture of acid and ore fed to the reaction stage is heated at the front section of the stage.
  • the unreacted acid is evaporated from the acid-treated laterite. Evaporation takes place for example by heating the acid-treated laterite to the boiling point of the acid at normal pressure or using negative pressure.
  • the slurry exiting the water leaching stage is routed to a neutralisation stage, in which iron is precipitated by neutralising the slurry.
  • the slurry formed in water leaching is routed directly to ion exchange treatment, where the ion exchange resin is selective with regard to nickel and cobalt.
  • the waste slurry is routed to neutralisation for the precipitation of other metals.
  • the slurry formed in water leaching is routed to a cementation stage, where the valuable metals nickel and cobalt are cemented from the solution by means of iron powder. After the cementation and magnetic separation of valuable metals, the waste slurry is routed to neutralisation for the precipitation of other metals.
  • a neutralising agent is fed to the water leaching stage of the acid-treated material in order to precipitate the iron as a hydroxide, while the nickel and cobalt remain as water-soluble salts.
  • the neutralising agent is preferably lime and/or lime milk. After the leaching and iron precipitation stage, the precipitation of nickel and cobalt as hydroxide, sulphide, or carbonate is performed.
  • the mineral acid used is sulphuric acid
  • at least some of the acid-treated material is routed to thermal treatment, in which the iron sulphates are broken down into sulphur oxides and hematite.
  • the sulphur oxides are routed to the sulphuric acid plant to manufacture sulphuric acid, which is used for the acid treatment of laterite ore.
  • the material exiting thermal treatment containing hematite and water-soluble metal salts is routed to water leaching.
  • Figure 1 is a flow diagram of one embodiment of the invention.
  • FIGS. 2 - 5 are flow diagrams of other embodiments of the invention.
  • nickel laterites such as limonite, saprolite and nontronite or their compounds
  • concentrated mineral acid 70-98% so that the metals contained in the laterites form water-soluble salts of the mineral acid.
  • the metals contained in laterites are mostly nickel, cobalt, manganese, magnesium, aluminium, chromium and iron, of which nickel and cobalt are mainly considered to be the valuable metals.
  • Mineral acid refers mostly to sulphuric acid, hydrochloric acid or nitric acid, or a mixture of them.
  • liquid-solid separation has proved problematic in the atmospheric leaching of laterites.
  • the object of this invention is to achieve a method whereby the liquid/solid separation properties are good and thus it is possible to make the method an economic one.
  • the laterite is dry and sufficiently fine in particle size.
  • normal concentrate coarseness is enough for the laterite grinding degree, in other words the method does not require actual fine grinding.
  • the grain size is too coarse, the acid will not be able to penetrate inside the laterite grains, and as a result the formation of water-soluble salts, i.e. in the case of sulphuric acid sulphation, will be slow and will often remain incomplete.
  • moist laterite causes the activation of the sulphation reaction due to the solvation heat of the acid, which hampers the mixing together of the laterite and acid.
  • laterite ore is crushed to a grain size of 90 % under 10 mm.
  • the crushed ore is dried in a drying stage 1 utilising the water vapour generated in the sulphation reactions of the later reaction stage 4.
  • Dried laterite ore is routed to a dry grinding stage 2, where it is ground typically to a grain size of 90 % under 500 ⁇ m, preferably 90 % under 150 ⁇ m, using for instance a ball mill.
  • the laterite is routed to the acid treatment mixing stage 3, in which laterite ore and concentrated sulphuric acid are mixed with each other homogenously in some suitable equipment, such as for example a screw mixer, drum-type reactor, or some other kind of reactor. Acid is added in such a quantity that it is at least in stoichiometric ratio with regard to the metals in the laterite.
  • the temperature of the mixing stage is adjusted preferably to be below 100 0 C, so that the generated mixture does not harden and complicate further processing.
  • the purpose is to operate at a temperature at which sulphation does not yet take place.
  • the mixture is transferred to the reaction stage 4, in which the process temperature is raised first by means of external heating, after which the reactions between the laterite and acid begin to produce heat, and the process becomes largely autothermal.
  • the reaction stage occurs at ambient pressure and at a temperature of around 150-300° C. Heat may be added in the early stages of the process, e.g. by means of internal or external heating such as burners, in order to optimize the desired chemical reactions.
  • the process temperature may be a maximum of the boiling point of the acid, i.e. the operating temperature is below 339 0 C when sulphuric acid is concerned. Laterites typically include a lot of crystalline water and in addition, for example in sulphation reactions, a considerable amount of water is generated, which evaporates during heating.
  • the water vapour generated is recovered and utilised in laterite drying, as mentioned above.
  • water is added to the sulphation stage, but in the method accordant with this invention water is not added to the reaction stage. Both the mixing and reaction stage like the later leaching stage all occur at ambient pressure, in other words the system is atmospheric.
  • the material exiting acid treatment is solid and powdery in nature and is thus easy to process.
  • the solid fine material that contains the water-soluble salts of the acid is routed to the actual leaching stage 6, where water is routed to the solids.
  • the water leaching stage takes place in atmospheric conditions, i.e. at a temperature of 80 - 105 0 C and under ambient pressure.
  • the duration of the leaching stage depends on the grain size and composition of the laterite ore and is typically between 1 - 2 h.
  • all the sulphated metals contained in the laterite dissolve.
  • iron which is mostly trivalent, is separated from the metal-containing solution, by neutralising the solution in the neutralisation stage 7 with some suitable neutralising agent, so that the iron is precipitated but the nickel and cobalt salts remain water-soluble.
  • Preferred neutralising agents are limestone and/or limestone milk, whereby iron is precipitated as hydroxide. If it is wished to precipitate the iron as jarosite for example, other known precipitation agents are used in addition, such as sodium sulphate. Aluminium and the majority of the chromium are precipitated at the same time as the iron, whereby they enter the leach residue.
  • solids separation 8 is performed, in which the solids are separated from the liquid in typical separation ways, such as thickening and/or filtering, whereby the substances left in the solids, such as iron and silicates, are made to separate from the solution.
  • All the dissolved metals and particularly the solution containing valuable metals (PLS) is routed to the following stage for further treatment of the solution, where the dissolved ions are made to separate effectively and routed to the next process stages (not shown in detail in the chart).
  • PLS valuable metals
  • the removal of magnesium and manganese from the solution occurs at a high pH value, typically in the range of 9 -1 1 , using lime or a corresponding neutralising agent, whereby Mg and Mn are precipitated, mostly as hydroxides.
  • Figure 2 presents a second alternative embodiment of the invention for treating the residue that exits the water leaching stage 6.
  • the initial stages of the method are performed as described above. It is typical of this alternative that no liquid-solid separation is carried out at all after leaching, but the entire amount of slurry is routed to ion exchange treatment 9, in which valuable metals such as nickel and cobalt are recovered by means of the ion exchange resin directly from the slurry in accordance with what is known as the resin-in-pulp concept. After adsorption to the ion exchange resin, the resin is routed to the elution stage 10, where it is eluted with acid or an equivalent eluent and the valuable metal containing elution solution is routed to further treatment.
  • the resin is recycled back to the ion exchange stage 9.
  • the valuable metals are recovered from the elution solution for example by means of sulphide precipitation or liquid-liquid extraction, which are not described in detail.
  • the waste sludge exiting the ion exchange stage, containing iron and silicates, is routed to the neutralisation stages, in which iron, aluminium, manganese and magnesium, etc. are precipitated out of the solution.
  • Figure 3 shows a third alternative embodiment of the invention for treating the residue that exits water leaching 6.
  • the desired valuable metals such as nickel and cobalt
  • the slurry is not subjected to liquid-solid separation before valuable metal recovery in this case either.
  • the iron powder is separated from the slurry in a magnetic separation stage 12, for instance by means of a weak magnetic separator.
  • the mixture of iron powder and valuable metals is routed to a valuable metals leaching stage 13, in which the mixture is leached with acid and the solution is routed to further treatment to recover nickel and cobalt.
  • the waste sludge generated during the cementation of valuable metals which contains iron and silicates, is routed to the neutralisation stages, where iron, aluminium, manganese and magnesium, etc. are precipitated out of the solution.
  • the process accordant with the invention as disclosed in Figure 1 may also be modified in accordance with Figure 4, where it is shown that the water leaching stage 6 following acid treatment may also be carried out as a combined leaching stage 14, in which the acid-treated laterite ore is treated directly with a neutralising solution by adding the amount of limestone required for the aqueous phase.
  • the leaching of the water- soluble salts of the laterite, solution neutralisation, and the precipitation of the resulting iron all take place in the same stage.
  • only one solids and liquid separation stage 15 is required, in which both the leach residue and the gypsum and iron precipitate that are formed are separated from the solution.
  • the PLS solution containing nickel and cobalt goes to further treatment, for instance sulphide precipitation 16, and the end solution goes to a magnesium and manganese removal stage 17, in which the metals are precipitated from the solution by neutralisation.
  • the end solution is mostly water, which can be recycled back to the leaching stage (not shown in detail in the drawing).
  • FIG. 5 Yet another alternative embodiment of the invention for the further treatment of acid-treated laterite is disclosed in Figure 5.
  • the material exiting the mixing and reaction stages is fed after the acid recovery stage 5 either wholly or partially into a thermal treatment stage 18, which may take place for example in a drum or fluidised bed furnace.
  • Thermal energy is required in the thermal treatment stage, which is obtained for example by burning coal.
  • the acid is sulphuric acid
  • the iron in the laterite reacts to form iron sulphate, which is now broken down in thermal treatment into hematite and sulphur oxide.
  • Sulphur oxide gases SO 2 and SO 3
  • the hematite-containing material is routed to water leaching 6, in which nickel, cobalt and magnesium dissolve and are recovered by the methods described above.
  • the free acid remaining in the acid-treated laterite may be evaporated as described earlier before thermal decomposition.
  • the advantages of the procedure are a waste material that is easy to treat and reduced acid costs.
  • EXAMPLES The operating method is illustrated with the following examples.
  • the laterite ore used in the examples is nontronitic and its composition is presented in Table 1.
  • the ore was dried and crushed to a size of 100% below 1 mm before the tests.
  • Dry laterite was ground with a ball mill to a grain size of 95% below 105 ⁇ m.
  • the acid used was sulphuric acid. Sulphuric acid and laterite were mixed together in the mixing stage at a mass ratio of 1 :1. 348 g of the mixture was fed into the reaction stage rotary kiln, where the temperature in the central point of the kiln was 25O 0 C. The residence time of the mixture in the hot part of the kiln was about 30 min. The weight of the rotary kiln product was 296.3 g.
  • the composition of the rotary kiln product is presented in Table 2. Table 2. Composition Ot the rotary kiln product.
  • the importance of grinding for nickel yield in leaching is illustrated by the following example.
  • the test was carried out with unground ore, which had been crushed to a grain size of 100% below 1 mm. Sulphuric acid and laterite were mixed together in the mixing stage at a mass ratio of 1 :1.
  • the reaction stage took place in a rotary kiln, into which 400g of the mixture was fed.
  • the temperature in the central point of the kiln was 25O 0 C.
  • the residence time of the mixture in the hot part of the kiln was about 27 min.
  • the weight of the rotary kiln product was 279.1 g.
  • the composition of the rotary kiln product is presented in Table 5.

Abstract

The invention relates to a method for processing laterite ores so that the metals contained in the laterites are turned into water-soluble form for the recovery of valuable metals, such as nickel and cobalt. Different types of nickel laterites are processed at the same time without being separated according to their iron and/or magnesium content. When laterites are pretreated with concentrated mineral acid so that the metals contained in the laterites react to form water-soluble salts, the silicates contained in the laterites are partially decomposed and postleaching liquid-solid separation becomes easier than it was previously. According to the method, the water vapour generated in the ore and acid reaction stage is utilised in ore drying and the unreacted mineral acid is recycled to the front end of the process.

Description

METHOD FOR TREATING NICKEL LATERITE ORE
FIELD OF THE INVENTION
The invention relates to a method for processing laterite ores so that the metals contained in the laterites are turned into water-soluble form for the recovery of valuable metals, such as nickel and cobalt. Different types of nickel laterites are processed at the same time without being separated according to their iron and/or magnesium content. When laterites are pre- treated with concentrated mineral acid so that the metals contained in the laterites react to form water-soluble salts, the silicates contained in the laterites are partially decomposed and post-leaching liquid-solid separation becomes easier than it was previously. According to the method, the water vapour generated in the ore and acid reaction stage is utilised in ore drying and the unreacted mineral acid is recycled to the front end of the process.
BACKGROUND OF THE INVENTION
A method is disclosed in US patent 4125588 in which nickel laterite ore is pre-treated with concentrated acid before nickel leaching. In the method described, laterite ore is dried so that the moisture content of the ore is less than 1 %. The dried ore is ground to a particle size range of 65 - 100 mesh. The ground ore is mixed into concentrated acid at a mass ratio of about 1 :1. The metal sulphation reactions are initiated by adding water to the mixture containing laterite and acid in a ratio of 3 - 40% of the mass of the laterite. The sulphated metals are leached into water.
US patent application 2006/0002835 describes a method in which laterite leaching takes place in two stages. In the first stage, the laterite ore is mixed with concentrated sulphuric acid. In the second stage, the ore/acid mixture is slurried with water and then leached, so that the nickel and cobalt dissolve. It is characteristic of the method that the amount of sulphuric acid used in pre- treatment is stoichimetric with regard to the non-ferrous metals in the ore, but not the iron. According to example 5, the acid/ore ratio is 0.65. A little excess acid is advantageous so that a small amount of iron also dissolves, because this achieves the maximal dissolution of nickel and cobalt. The leaching stage of the method is implemented either at a temperature of 95 - 105° C or in an autoclave at a temperature which is a maximum of 150° C and in which the pressure corresponds to the saturated steam pressure. To improve the dissolution of cobalt, some suitable reductant is added to the leaching stage, such as sulphur dioxide. Owing to the sub-stoichiometric acid usage in the early stage of leaching treatment, only a part of the laterite is dissolved. Remarkably long retention times have to be used in the method for the leaching-precipitation stage. The effectiveness of the method requires that there is a significant proportion of easily-soluble saprolitic laterite in the laterite, so that it can function in the ferric iron precipitation range.
Other methods in the prior art for recovering nickel and other valuable metals from laterites are fairly comprehensively described in US patent application 2006/0002835.
Since an atmospheric leaching process generates a very fine and gel-like leach residue, the separation and washing of solids and solution are especially demanding and pose some of the greatest challenges for the method. In spite of the many process patents based on atmospheric leaching, this problem has not been paid sufficient attention. The difficulty of solid-liquid separation also varies to a considerable extent depending on the type of laterite.
PURPOSE OF THE INVENTION
The purpose of the invention presented here is to eliminate the problems that have arisen in earlier methods and to achieve a method for leaching laterites in a way that in particular solids separation from the solution does not cause problems. The invention also aims to improve the economics of the laterite process by utilising the heat generated in the reactions and recovering the acid that is not consumed in the reactions. SUMMARY OF THE INVENTION
The invention relates to a method for processing nickel laterite ore to facilitate the recovery of nickel and cobalt and liquid-solid separation. In accordance with the method a) crushed laterite ore is dried using the steam from a later process stage, b) the dried ore is subjected to dry grinding, c) the fine ore is routed to an acid treatment mixing stage that takes place using concentrated mineral acid, in which the ore is mixed homogenously into the acid, the amount of which is at least stoichiometric with regard to the metals contained in the ore, d) the mixture of ore and acid is fed to the acid treatment reaction stage, which takes place at ambient pressure and a temperature of between 15O0C and the boiling point of the acid, whereby the acid and ore react with each other; the water vapour formed in the reaction is recovered and recycled for use in ore drying, e) the acid-treated laterite ore is routed to acid recovery, where the acid that remained unreacted in the reactions of the acid treatment mixing stage is recovered by evaporation, after which the acid is cooled and recycled back to the mixing stage, f) the acid-treated laterite ore exiting acid recovery is routed to the metals leaching stage, which takes place with water.
It is typical of the method accordant with the invention that the particle size of the crushed laterite ore is in the region of 90% below 10 mm. In the dry grinding stage, it is advantageous to grind the ore to a particle size of 90 % below 500 μm, preferably 90% below 150 μm.
In accordance with the method, the acid and laterite are mixed into a homogenous mixture at a temperature where the metals do not yet substantially react with the acid. In the acid treatment accordant with the method, which includes the acid and ore mixing stage and the reaction stage, the metals of the laterite ore are made to form water-soluble salts of the mineral acid. The mineral acid is at least one of the following: sulphuric acid, nitric acid, or hydrochloric acid, or a mixture of at least two of them. The concentration of the mineral acid is preferably in the order of 70 - 98 wt.%. When the mineral acid is sulphuric acid, the reaction stage is sulphation.
It is typical of the method accordant with the invention that the mixture of acid and ore fed to the reaction stage is heated at the front section of the stage. In the method it is typical of the acid recovery stage that the unreacted acid is evaporated from the acid-treated laterite. Evaporation takes place for example by heating the acid-treated laterite to the boiling point of the acid at normal pressure or using negative pressure.
According to one embodiment of the invention, the slurry exiting the water leaching stage is routed to a neutralisation stage, in which iron is precipitated by neutralising the slurry.
According to another embodiment of the invention, the slurry formed in water leaching is routed directly to ion exchange treatment, where the ion exchange resin is selective with regard to nickel and cobalt. After ion exchange treatment, the waste slurry is routed to neutralisation for the precipitation of other metals.
According to a third embodiment of the invention, the slurry formed in water leaching is routed to a cementation stage, where the valuable metals nickel and cobalt are cemented from the solution by means of iron powder. After the cementation and magnetic separation of valuable metals, the waste slurry is routed to neutralisation for the precipitation of other metals. According to a fourth embodiment of the invention, a neutralising agent is fed to the water leaching stage of the acid-treated material in order to precipitate the iron as a hydroxide, while the nickel and cobalt remain as water-soluble salts. The neutralising agent is preferably lime and/or lime milk. After the leaching and iron precipitation stage, the precipitation of nickel and cobalt as hydroxide, sulphide, or carbonate is performed.
According to yet another embodiment of the invention, the mineral acid used is sulphuric acid, and at least some of the acid-treated material is routed to thermal treatment, in which the iron sulphates are broken down into sulphur oxides and hematite. The sulphur oxides are routed to the sulphuric acid plant to manufacture sulphuric acid, which is used for the acid treatment of laterite ore. The material exiting thermal treatment containing hematite and water-soluble metal salts is routed to water leaching.
LIST OF DRAWINGS
Figure 1 is a flow diagram of one embodiment of the invention, and
Figures 2 - 5 are flow diagrams of other embodiments of the invention.
DETAILED DESCRIPTION OF THE INVENTION
In the method accordant with the invention, all different types of nickel laterites, such as limonite, saprolite and nontronite or their compounds, are treated together with concentrated mineral acid (70-98%) so that the metals contained in the laterites form water-soluble salts of the mineral acid. Even if the laterite to be treated were to be formed of several different types of laterite, in the method accordant with the invention they would nevertheless not be separated from each other. The metals contained in laterites are mostly nickel, cobalt, manganese, magnesium, aluminium, chromium and iron, of which nickel and cobalt are mainly considered to be the valuable metals. Mineral acid refers mostly to sulphuric acid, hydrochloric acid or nitric acid, or a mixture of them. Below for the sake of simplicity we mention sulphuric acid, but the method is also applicable to other mineral acids. Likewise for simplicity the terms sulphation and sulphated laterite are used, but these terms are not meant to limit the use of the method accordant with the invention by means of other acids.
Generally, liquid-solid separation has proved problematic in the atmospheric leaching of laterites. The object of this invention is to achieve a method whereby the liquid/solid separation properties are good and thus it is possible to make the method an economic one.
It is particularly important for the process that the mixing of acid and laterite can be performed in optimal conditions. In this case it is essential that the laterite is dry and sufficiently fine in particle size. However, normal concentrate coarseness is enough for the laterite grinding degree, in other words the method does not require actual fine grinding. If the grain size is too coarse, the acid will not be able to penetrate inside the laterite grains, and as a result the formation of water-soluble salts, i.e. in the case of sulphuric acid sulphation, will be slow and will often remain incomplete. On the other hand, the use of moist laterite causes the activation of the sulphation reaction due to the solvation heat of the acid, which hampers the mixing together of the laterite and acid.
The method accordant with the invention is described in the appended Figure 1 with a flow chart. In accordance with the method, laterite ore is crushed to a grain size of 90 % under 10 mm. The crushed ore is dried in a drying stage 1 utilising the water vapour generated in the sulphation reactions of the later reaction stage 4. Dried laterite ore is routed to a dry grinding stage 2, where it is ground typically to a grain size of 90 % under 500 μm, preferably 90 % under 150 μm, using for instance a ball mill.
After drying and grinding, the laterite is routed to the acid treatment mixing stage 3, in which laterite ore and concentrated sulphuric acid are mixed with each other homogenously in some suitable equipment, such as for example a screw mixer, drum-type reactor, or some other kind of reactor. Acid is added in such a quantity that it is at least in stoichiometric ratio with regard to the metals in the laterite. The temperature of the mixing stage is adjusted preferably to be below 1000C, so that the generated mixture does not harden and complicate further processing. The purpose is to operate at a temperature at which sulphation does not yet take place.
After the mixing stage, the mixture is transferred to the reaction stage 4, in which the process temperature is raised first by means of external heating, after which the reactions between the laterite and acid begin to produce heat, and the process becomes largely autothermal. The reaction stage occurs at ambient pressure and at a temperature of around 150-300° C. Heat may be added in the early stages of the process, e.g. by means of internal or external heating such as burners, in order to optimize the desired chemical reactions. The process temperature may be a maximum of the boiling point of the acid, i.e. the operating temperature is below 3390C when sulphuric acid is concerned. Laterites typically include a lot of crystalline water and in addition, for example in sulphation reactions, a considerable amount of water is generated, which evaporates during heating. The water vapour generated is recovered and utilised in laterite drying, as mentioned above. In the method accordant with US patent 4,125,588, water is added to the sulphation stage, but in the method accordant with this invention water is not added to the reaction stage. Both the mixing and reaction stage like the later leaching stage all occur at ambient pressure, in other words the system is atmospheric. The material exiting acid treatment is solid and powdery in nature and is thus easy to process.
The use of excess acid with regard to the metals contained in the laterite improves the transformation of the metals into water-soluble salts (sulphation) and thus improves the recovery degree of the metals. The drawback in this case is the increase in costs of acid and neutralisation. Acid consumption in the process is reduced by using the recycling of unreacted residual acid. Residual acid is evaporated from the acid-treated laterite in the acid recovery stage 5. Evaporation takes place for example by heating acid- treated laterite to the boiling point of the acid, 3390C. Another alternative is to use negative pressure, so that the boiling point of the acid is lower. After evaporation the acid is allowed to cool and condense into liquid, which is recycled back to the mixing stage 3.
When laterite reacts for example with sulphuric acid, it can be described by the following simplified reaction equations:
2 FeOOH + 3 H2SO4 => Fe2(SO4)3 + 4H2O (1 )
NiO + H2SO4 => NiSO4 + H2O (2)
2 CoOOH + 3 H2SO4 => Co2(SO4)3 + 4H2O (3)
AI2O3 + 3 H2SO4 => AI2(SO4)3 + 3 H2O (4) MgO + H2SO4 => MgSO4 + H2O (5)
M2O + H2SO4 => M2SO4 + H2O (M = Na, K) (6)
In the method accordant with the invention, phenomena are utilised according to which the silicates contained in the laterites are partially dehydrated in concentrated acid treatment and possibly also in stages that take place at higher temperatures. At the same time their crystal structure partially changes in a way that facilitates the subsequent liquid-solid separation.
After the acid treatment and acid recovery stage accordant with Figure 1 , the solid fine material that contains the water-soluble salts of the acid is routed to the actual leaching stage 6, where water is routed to the solids. The water leaching stage takes place in atmospheric conditions, i.e. at a temperature of 80 - 1050C and under ambient pressure. The duration of the leaching stage depends on the grain size and composition of the laterite ore and is typically between 1 - 2 h. In the leaching stage 6 of acid-treated laterite ore, all the sulphated metals contained in the laterite dissolve. In the alternative accordant with Figure 1 , iron, which is mostly trivalent, is separated from the metal-containing solution, by neutralising the solution in the neutralisation stage 7 with some suitable neutralising agent, so that the iron is precipitated but the nickel and cobalt salts remain water-soluble. Preferred neutralising agents are limestone and/or limestone milk, whereby iron is precipitated as hydroxide. If it is wished to precipitate the iron as jarosite for example, other known precipitation agents are used in addition, such as sodium sulphate. Aluminium and the majority of the chromium are precipitated at the same time as the iron, whereby they enter the leach residue.
After the water leaching stage 6 and neutralisation 7, solids separation 8 is performed, in which the solids are separated from the liquid in typical separation ways, such as thickening and/or filtering, whereby the substances left in the solids, such as iron and silicates, are made to separate from the solution. All the dissolved metals and particularly the solution containing valuable metals (PLS) is routed to the following stage for further treatment of the solution, where the dissolved ions are made to separate effectively and routed to the next process stages (not shown in detail in the chart). As stated above, in the treatment of laterites in accordance with the invention, the structure of the silicates is changed so that solution-solids separation does not cause gelation and is therefore easy to perform. It is thus typical of the method that, owing to the properties of the leach residue, the thickening and scrubbing of the solution can be performed with considerably smaller equipment than for instance after direct acid leaching. At the same time, the proportion of valuable metals remaining in the final waste can also be made significantly smaller than in known methods.
The recovery of desired valuable metals such as nickel and cobalt from the solution occurs by known methods, by precipitating them after iron precipitation as hydroxide, sulphide, or carbonate, nor are they shown in detail in the chart. Typical chemicals are used in precipitation such as e.g. hydrogen sulphide, sodium hydrogen sulphate, lime, magnesium hydroxide, or sodium carbonate.
The removal of magnesium and manganese from the solution occurs at a high pH value, typically in the range of 9 -1 1 , using lime or a corresponding neutralising agent, whereby Mg and Mn are precipitated, mostly as hydroxides.
Figure 2 presents a second alternative embodiment of the invention for treating the residue that exits the water leaching stage 6. The initial stages of the method are performed as described above. It is typical of this alternative that no liquid-solid separation is carried out at all after leaching, but the entire amount of slurry is routed to ion exchange treatment 9, in which valuable metals such as nickel and cobalt are recovered by means of the ion exchange resin directly from the slurry in accordance with what is known as the resin-in-pulp concept. After adsorption to the ion exchange resin, the resin is routed to the elution stage 10, where it is eluted with acid or an equivalent eluent and the valuable metal containing elution solution is routed to further treatment. The resin is recycled back to the ion exchange stage 9. The valuable metals are recovered from the elution solution for example by means of sulphide precipitation or liquid-liquid extraction, which are not described in detail. The waste sludge exiting the ion exchange stage, containing iron and silicates, is routed to the neutralisation stages, in which iron, aluminium, manganese and magnesium, etc. are precipitated out of the solution.
Figure 3 shows a third alternative embodiment of the invention for treating the residue that exits water leaching 6. The desired valuable metals, such as nickel and cobalt, are recovered from the slurry exiting the leaching stage 6 in a cementation stage 1 1 , by cementing them with iron powder of suitable coarseness, whereby the valuable metals remain in the iron powder. The slurry is not subjected to liquid-solid separation before valuable metal recovery in this case either. The iron powder is separated from the slurry in a magnetic separation stage 12, for instance by means of a weak magnetic separator. The mixture of iron powder and valuable metals is routed to a valuable metals leaching stage 13, in which the mixture is leached with acid and the solution is routed to further treatment to recover nickel and cobalt. The waste sludge generated during the cementation of valuable metals, which contains iron and silicates, is routed to the neutralisation stages, where iron, aluminium, manganese and magnesium, etc. are precipitated out of the solution.
The process accordant with the invention as disclosed in Figure 1 may also be modified in accordance with Figure 4, where it is shown that the water leaching stage 6 following acid treatment may also be carried out as a combined leaching stage 14, in which the acid-treated laterite ore is treated directly with a neutralising solution by adding the amount of limestone required for the aqueous phase. In this case, the leaching of the water- soluble salts of the laterite, solution neutralisation, and the precipitation of the resulting iron all take place in the same stage. In this alternative only one solids and liquid separation stage 15 is required, in which both the leach residue and the gypsum and iron precipitate that are formed are separated from the solution. The PLS solution containing nickel and cobalt goes to further treatment, for instance sulphide precipitation 16, and the end solution goes to a magnesium and manganese removal stage 17, in which the metals are precipitated from the solution by neutralisation. The end solution is mostly water, which can be recycled back to the leaching stage (not shown in detail in the drawing).
Yet another alternative embodiment of the invention for the further treatment of acid-treated laterite is disclosed in Figure 5. The material exiting the mixing and reaction stages is fed after the acid recovery stage 5 either wholly or partially into a thermal treatment stage 18, which may take place for example in a drum or fluidised bed furnace. Thermal energy is required in the thermal treatment stage, which is obtained for example by burning coal. When the acid is sulphuric acid, in the acid treatment the iron in the laterite reacts to form iron sulphate, which is now broken down in thermal treatment into hematite and sulphur oxide. Sulphur oxide gases (SO2 and SO3) are routed on to the acid plant 19, from where the acid generated is recycled back to the acid treatment stage 3. The hematite-containing material is routed to water leaching 6, in which nickel, cobalt and magnesium dissolve and are recovered by the methods described above. The free acid remaining in the acid-treated laterite may be evaporated as described earlier before thermal decomposition. The advantages of the procedure are a waste material that is easy to treat and reduced acid costs.
EXAMPLES The operating method is illustrated with the following examples. The laterite ore used in the examples is nontronitic and its composition is presented in Table 1. The ore was dried and crushed to a size of 100% below 1 mm before the tests.
Table 1. Composition of nickel laterite.
Al Co Fe Mg Ni
3.6 0.036 14.6 5.1 0.72
Example 1.
Dry laterite was ground with a ball mill to a grain size of 95% below 105 μm. The acid used was sulphuric acid. Sulphuric acid and laterite were mixed together in the mixing stage at a mass ratio of 1 :1. 348 g of the mixture was fed into the reaction stage rotary kiln, where the temperature in the central point of the kiln was 25O0C. The residence time of the mixture in the hot part of the kiln was about 30 min. The weight of the rotary kiln product was 296.3 g. The composition of the rotary kiln product is presented in Table 2. Table 2. Composition Ot the rotary kiln product.
Al Co Fe Mg Ni
% % % % %
1 .08 0.031 8 .82 2.74 0.48
After treatment in the rotary kiln, 294.8 g of the suphated laterite was leached with 1 litre of water in a titanium reactor in the leaching stage. The duration of the leaching test was 6 hours and the temperature of the slurry was 8O0C. The metal content of the solids during leaching is presented in Table 3. The weight of the leach residue was 140.2 g. The metal yields to solution are presented in Table 4.
Table 3. Metal content of solids during leaching.
Leaching time Al Co Fe Mg Ni h % % % % %
0 1 .08 0.031 8.82 2.74 0.48
1 0.27 <0.004 2.25 0.8 0.061
2 0.24 <0.004 2.04 0.75 0.056
4 0.26 <0.004 2.07 0.74 0.052
6 0.25 <0.004 1 .96 0.73 0.05
Table 4. Metal yields to solution.
Al Co Fe Mg Ni
% % % % %
89.0 >93.9 89.4 87.3 95.0
Example 2
The importance of grinding for nickel yield in leaching is illustrated by the following example. The test was carried out with unground ore, which had been crushed to a grain size of 100% below 1 mm. Sulphuric acid and laterite were mixed together in the mixing stage at a mass ratio of 1 :1. The reaction stage took place in a rotary kiln, into which 400g of the mixture was fed. The temperature in the central point of the kiln was 25O0C. The residence time of the mixture in the hot part of the kiln was about 27 min. The weight of the rotary kiln product was 279.1 g. The composition of the rotary kiln product is presented in Table 5.
Table 5. Composition Ot rotary kiln product.
Al Co Fe Mg Ni
% % % % %
1 .8 0.016 7.9 2.6 0.38
After treatment in the rotary kiln, 200 g of the sulphated laterite was leached with 1 litre of water in a 1 -litre titanium reactor in the leaching stage. The duration of the leaching test was 6 hours and the temperature of the slurry was 8O0C. The metal content of the solids during leaching is presented in Table 6. The weight of the leach residue was 66.2 g. The metal yields to solution are presented in Table 7.
Table 6. Metal contents of solids during leaching
Leaching time Al Co Fe Mg Ni h % % % % %
0 1 .8 0.016 7.9 2.6 0.38
1 1 .26 0.008 7.37 2.56 0.302
2 1 .25 0.005 7.34 2.55 0.289
4 1 .17 0.005 7.07 2.52 0.267
6 1 .12 0.004 7.03 2.59 0.255
Table 7. Metal yields to solution.
Al Co Fe Mg Ni
% % % % %
79.4 91 .7 70.5 67.0 77.8
Example 3
The importance of the mass ratio of acid and laterite for nickel yield in leaching is illustrated by the following example. Two more tests were carried out as in example 1 , apart from the fact that acid was used in ratios of 0.6 and 0.8 kg/kg laterite. The amounts of acid not consumed in the sulphation reactions of the treatment were determined after water leaching by titrating the acid remaining in the solution. Table 8 presents the effect of the acid to laterite ratio on metal yields and on the amount of unreacted acid. The table shows that metal yields clearly improve as the amount of acid increases in relation to the weight of laterite. In this case the amount of unreacted acid also increases. The result shows that for maximal nickel yield acid should be used in amounts that are at a mass ratio of about 1 :1 with regard to laterite. When acid is recycled in the way presented in this invention, substantial cost savings can be achieved in relation to acid consumption.
Table 8. Metal yields to solution and amounts of residual acid with different acid-laterite ratios.
Acid:laterite Al Co Fe Mg Ni Residual acid kg/kg % % % % % kg/kg laterite
0.6 91 .3 89.7 79.4 89.1 83.0 0.07
0.8 96.0 89.7 89.4 93.3 92.3 0.13
1 89.0 >93.9 89.4 87.3 95.0 0.28

Claims

PATENT CLAIMS
1. A method for treating nickel laterite ore in order to recover nickel and cobalt and facilitate liquid-solid separation, characterised in that in accordance with the method a) crushed laterite ore is subjected to drying (1 ) by means of steam from a later process stage, b) the dried ore is subjected to dry grinding (2), c) the fine ore is routed to an acid treatment mixing stage (3) that uses concentrated mineral acid, in which the ore is mixed homogenously into acid, which is quantitatively at least stoichiometric with regard to the metals contained in the ore, d) the mixture of ore and acid is fed into an acid treatment reaction stage (4), which occurs at ambient pressure and a temperature between 15O0C and the boiling point of the acid, so that the acid and ore react with each other; the water vapour formed in the reactions is recovered and recycled for use in ore drying (1 ), e) the acid-treated laterite ore is routed to acid recovery (5), where unreacted acid is recovered by evaporation, after which the acid is cooled and recycled back to the mixing stage (3), f) the acid-treated laterite ore exiting acid recovery is routed to a metals leaching stage (6), which takes place with water.
2. A method according to claim 1 , characterised in that the grain size of the crushed laterite ore is in the region of 90% below 10 mm.
3. A method according to claim 1 , characterised in that in the dry grinding stage (2) the ore is crushed to a grain size of 90 % below 500 μm.
4. A method according to claim 3, characterised in that in the dry grinding stage (2) the ore is crushed to a grain size of 90 % below 150 μm.
5. A method according to claim 1 , characterised in that the acid and laterite are mixed into a homogenous mixture at a temperature where the metals not yet react with the acid.
6. A method according to claim 1 , characterised in that the metals of the laterite ore are made to form water-soluble salts of the mineral acid in acid treatment (3,4).
7. A method according to claim 1 , characterised in that the mineral acid is at least one of the following: sulphuric acid, nitric acid or hydrochloric acid, or a mixture of at least two of these.
8. A method according to claim 1 , characterised in that the concentration of the mineral acid is 70 - 98%.
9. A method according to claims 1 and 7, characterised in that the mineral acid is sulphuric acid and the reaction stage (4) is sulphation.
10. A method according to claim 1 , characterised in that the mixture of acid and ore fed into the reaction stage (4) is heated in the initial part of the stage.
1 1. A method according to claim 1 , characterised in that the unreacted acid in the acid recovery stage (5) is evaporated by heating the acid- treated laterite at normal pressure to the boiling point of the acid.
12. A method according to claim 1 , characterised in that the unreacted acid in the acid recovery stage (5) is evaporated using negative pressure.
13. A method according to claim 1 , characterised in that the slurry exiting the water leaching stage (6) is routed to a neutralisation stage (7), in which iron is precipitated by neutralising the slurry.
14. A method according to claim 1 , characterised in that the slurry formed in water leaching (6) is routed directly to ion exchange treatment (9), in which the ion exchange resin is selective with regard to nickel and cobalt.
15. A method according to claim 14, characterised in that the waste sludge from ion exchange treatment (9) is routed to neutralisation to precipitate the other metals.
16. A method according to claim 1 , characterised in that the slurry formed in water leaching (6) is routed to a cementation stage (1 1 ), in which the valuable metals nickel and cobalt are cemented from solution by means of iron powder and are separated from the slurry by magnetic separation (12).
17. A method according to claim 16, characterised in that after the cementation (1 1 ) and magnetic separation (12) of the valuable metals, the waste sludge is routed to neutralisation to precipitate other metals.
18. A method according to claim 1 , characterised in that a neutralising agent is fed into the water leaching stage (14) of the acid-treated material to precipitate the iron as hydroxide while the nickel and cobalt remain as water-soluble salts.
19. A method according to claim 18, characterised in that the neutralising agent is limestone and/or lime milk.
20. A method according to claims 1 and 18, characterised in that after the leaching and iron precipitation stage (14), nickel and cobalt precipitation (16) is performed from the solution as hydroxide, sulphide or carbonate.
21. A method according to claim 1 , characterised in that the mineral acid used is sulphuric acid, and at least some of the acid-treated material is routed to thermal treatment (18), in which iron sulphates are broken down into sulphur oxides and hematite.
22. A method according to claim 21 , characterised in that the sulphur oxides are routed to a sulphuric acid plant (19) to manufacture sulphuric acid, which is used for the acid treatment of laterite ore.
23. A method according to claim 21 , characterised in that the material consisting of hematite and water-soluble metal salts exiting thermal treatment (18) is routed to water leaching (6).
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