FROM VARIOUS FEEDSTOCKS

FIELD OF THE INVENTION

[0001] The present invention relates generally to processes for the recovery of manganese and associated metal products in a pure form from various manganiferous feedstocks. In particular, the process is applied to materials containing calcium and potassium.

BACKGROUND OF THE INVENTION

[0002] Manganese is currently recovered from relatively high-grade concentrates by a variety of methods, both pyrometallurgical (ferro-manganese) and hydrometallurgical (electrolytic manganese metal and manganese dioxide, manganese carbonate and manganese sulphate). To date, concentrates have been produced wherein the impurity levels, especially calcium and potassium, are low. However, such feedstocks are dwindling, and increasingly it will be necessary to process materials with higher impurity levels. In particular, ores from N. America generally contain significant amounts of calcium, whereas those from N. Africa contain potassium.

[0003] Reductive leaching of manganese minerals, especially with sulphur dioxide gas, is well known. One such process is described by C.B. Ward in US Patent 7,951,282 B2, entitled Hydrometallurgical Processing of Manganese Containing Minerals, published May 31 , 2011. This process notes that although leaching with sulphur dioxide is well known, it results in solutions which contain significant levels of dithionate ions. In order to control the concentrations of dithionate to <5 g/L, care is taken to ensure that excess manganese dioxide mineral is maintained in the slurry. This, unfortunately results in reduced efficiency of leaching.

[0004] Further, in the process described, significant amounts of potassium are leached. In order to control and remove the potassium, rather than seeing the potassium as an opportunity for generating value-added products, the process adds a soluble source of iron in order to precipitate potassium jarosite. This is not only an imperfect process, since jarosite has a relatively high residual solubility, it is also very costly in terms of having to add iron, especially if the potassium levels are high. [0005] W. Zhang and C.Y. Cheng, in an article entitled Manganese Metallurgy Review, Part I: Leaching of Ores/Secondary Materials and Recovery of Electrolytic/Chemical Manganese Dioxide, published in Hydrometallurgy 89 (2007), pp. 137-159, have comprehensively reviewed many process for leaching manganese from ores and other feed materials. They conclude that using sulphur dioxide is the most effective method. However, unlike the process of Ward mentioned above, they do not regard the formation of dithionate as being problematic.

[0006] M.S. Bafghi, A. Zakeri, Z. Ghasemi and M. Adeli, in an article entitled Reductive Dissolution of Manganese Ore in Sulphuric Acid in the Presence of Iron Metal, published in Hydrometallurgy 90 (2008), pp. 207-212, have described a leaching process wherein metallic iron is used as the reductant rather sulphur dioxide. Based on the data presented, iron seems to be the more efficient reductant, but the requirements of approximately one tonne of iron per tonne of manganese render such a process entirely uneconomic.

[0007] N. Chow, A. Nacu, D. Warkentin, H. Teh and I. Askenov, in an article entitled New Developments in the Recovery of Manganese from Lower-Grade Resources, published in Minerals and Metallurgical Processing 29(1), February 2012, pp. 61-74, and in accompanying US Patent 8,460,631 B2, entitled Processing of Manganous Sulphate/Dithionate Liquors Derived from Manganese Resource Material by the same authors with the exception of H. Teh, published June 11, 2013, also discuss the problem of dithionates in the S0 ₂ leaching of manganese feedstocks. They discuss the various dithionates present, and how these can both be problematic and advantageous. The conclusion is that dithionates should not be an issue. It is noted in the article that the patent does not introduce any new technology, but rather is a novel combination of conventional processes.

[0008] Based on the above reviews, it appears that reductive leaching of manganese using sulphur dioxide is by far the most preferred method of dissolving manganese, but that there are potential problems with the formation of dithionates.

[0009] W. Zhang, P. Singh and D. Muir, in an article entitled Oxidative Precipitation of Manganese with S02/02 and Separation from Cobalt and Nickel, published in Hydrometallurgy, 63 (2002), pp. 127-135, and W. Zhang and C.Y. Cheng, in article entitled Manganese Metallurgy Review, Part II: Manganese Separation and Recovery from Solution, published in Hydrometallurgy 89 (2007), pp. 160-177, have reviewed processes for the recovery of manganese from solution. Oxidative precipitation is substantially the preferred method. Nominally, it would be expected that manganese dioxide, Mn0 ₂ would be recovered, but this is not the case, and mixed oxides of manganese, predominantly Mn ₂ 0 ₃ and Mn0 ₂ are recovered, thus making the product somewhat indeterminate and of little economic value.

[0010] It is postulated that mixed oxides are recovered due to the many various oxidation states manganese can exist in, and further because in precipitating the oxides, acid is released.

[0011] V. Menard and G.P. Demopoulos, in an article entitled Gas Transfer Kinetics and Redox Potential Considerations in Oxidative Precipitation of Manganese From an Industrial Zinc Sulphate Solution with S0 ₂ /0 ₂ published in Hydrometallurgy 89 (2007), pp. 357-368, also note that indeterminate manganese oxides are precipitated, and that they are likely hydrated. These authors also note that although apparently crystalline, the oxides unusually do not exhibit any discernible X-Ray Diffraction pattern, making characterisation of the precipitate difficult.

[0012] The focus of the process described in paragraph [0007] is on the recovery of initially manganese carbonate as a product in itself, and also as a precursor for the production of electrolytic manganese metal. The carbonate is precipitated in a conventional manner with soda ash, yielding a solution of sodium sulphate, which is then crystallised to the decahydrate (Na ₂ SO _4* l0H ₂ O), which in turn is calcined to the anhydrous salt. The reason for this is to be able to preserve the water balance, but evaporating so much water is very expensive.

[0013] A paper by David Dreisinger, Eric Norton, Ken Baxter, Michael Holmes and Ron Molnar, entitled The Hydrometallurgical Treatment of Baja Mining’s El Boleo Project for Cu, Co, Zn and Mn Recovery, presented at ALTA Copper 2008, Perth, WA, May 2008, also describes a process of reductive leaching with S0 ₂ and manganese carbonate recovery by precipitation with soda ash. This process mentions the presence of calcium, and that sub-stoichiometric addition of soda ash is necessary in order to prevent the co-precipitation calcium carbonate, which would contaminate the manganese carbonate product. This results in not only all of the manganese not being recovered, but also an appreciable recycle. [0014] It is thus clear that whilst preferred methods of leaching and precipitation are generally well-known, employing S0 ₂ and SO2/O2 mixtures and/or soda ash precipitation, these processes have difficulties that have not hitherto been overcome. Further, no attempt has been made to develop any processes for the economic recovery of perceived nuisance elements such as potassium and calcium, but rather to employ methods, sometimes very expensive, to affect their removal and disposal. There is a need for a process or processes that not only overcome the deficiencies of existing processes, but one which also recovers hitherto nuisance elements such as potassium and/or calcium as value-added products.

SUMMARY OF THE INVENTION

[0015] In a first aspect of the invention, there is provided a process for separating Ca from a Ca- and Mn-containing solid feed, the process including: leaching the solid feed stock with an aqueous leachant, the aqueous leachant including an acid to convert Ca in the solid feed to a soluble Ca salt and form a Ca-rich leachate and a Ca-lean Mn-containing solid residue; and separating the Ca-rich leachate and the Ca-lean Mn-containing solid residue.

[0016] In an embodiment, the process further includes treating the Ca-rich leachate with a precipitant to form a Ca-containing precipitate and a Ca-lean leachate. In one form, the process further includes separating the Ca-containing precipitate and the Ca-lean leachate. Preferably, the precipitant is selected from the group consisting of SO2 gas, or a soluble sulphite, and the precipitate is CaS0 ₃ .

[0017] In forms of the invention where the precipitant is SO2 gas and the precipitate is CaSCh, the process preferably further includes calcining the Ca-containing precipitate to form CaO and SO2 gas; and recycling the SO2 gas as a feed (or component thereof) to the Ca-precipitation step. It is preferred that the step of calcining the Ca-containing precipitate is carried out at a temperature of from about 500°C up to about l000°C. More preferably, the temperature is from about 600°C. Even more preferably the temperature is from about 700°C. Most preferably, the temperature is from about 750°C. Additionally or alternatively, it is preferred that the temperature is up to about 950°C. More preferably, the temperature is up to about 900°C. Most preferably, the temperature is up to about 850°C.

[0018] In an embodiment, the acid is selected from the group consisting of: HC1, HNO3, and organic acids, such as formic acid, and acetic acid.

[0019] In an embodiment, the step of leaching the solid feed stock is carried out at a temperature of from ambient up to l00°C. Preferably, the temperature is from about 40°C. More preferably the temperature is from about 45 °C. Most preferably, the temperature is from about 50°C. Additionally or alternatively, it is preferred that the temperature is up to about 90°C. More preferably, the temperature is up to about 80°C. Even more preferably, the temperature is up to about 70°C. Most preferably, the temperature is up to about 60°C.

[0020] In a second aspect of the invention, there is provided a method of separating Mn from a Ca-lean Mn-containing solid residue from the process of the first aspect, the process including: leaching the Ca-lean Mn-containing solid residue with sulphuric and a reductant to reduce Mn in the Ca-lean Mn-containing solid residue and form a MnSCU containing leachate and a Mn- lean solid residue; separating the nSCU containing leachate and the Mn-lean solid residue; contacting the MnSCU containing leachate with an alcohol to form a MnSCU precipitate and a Mn-lean liquor; and separating the nSCU precipitate and the Mn-lean liquor.

[0021] In an embodiment, the reductant is a reducing sulphur species. Preferably, the reducing sulphur species is SO2.

[0022] In a third aspect of the invention, there is provided a process for separating Mn from an Mn-containing solid feed, the process including: leaching the Mn-containing feed with sulphuric acid and a reductant to reduce Mn in the solid feed and form a nSCU containing leachate and a Mn-lean solid residue; separating the MnSOr containing leachate and the Mn-lean solid residue; contacting the MnSCU containing leachate with an alcohol to form an nSCU precipitate and a Mn-lean liquor; and separating the nSCU precipitate and the Mn-lean liquor.

[0023] In an embodiment, the reductant is a reducing sulphur species. Preferably, the reducing sulphur species is S0 ₂ .

[0024] In an embodiment of the third aspect, the Mn-containing solid feed is produced according to the first aspect of the invention.

[0025] In an embodiment of the second or third aspects, the process further includes: heating the Mn-lean liquor to separate the alcohol from the Mn-lean liquor and recycling the alcohol for use in the step of contacting the nSCU containing leachate with the alcohol.

[0026] In one form of the above embodiment, the step of heating the Mn-lean liquor to separate the alcohol includes distilling the alcohol from the Mn-lean liquor at a temperature of from about 60°C to about l00°C. Preferably, the temperature is from about 65°C. More preferably the temperature is from about 70°C. Most preferably, the temperature is from about 75°C. Additionally or alternatively, it is preferred that the temperature is up to about 95°C. More preferably, the temperature is up to about 90°C. Most preferably, the temperature is up to about 85°C. A particularly preferred temperature is about 80°C.

[0027] In one form of the above embodiment, the alcohol is selected from the group consisting of alcohol compounds having 1, 2, or 3 carbon atoms. Preferably, the alcohol is selected from the group consisting of methanol, ethanol, and 1 -propanol. Most preferably, the alcohol is methanol.

[0028] In embodiments of the second and third aspects, where K is additionally present in the Mn- containing feed, the reducing sulphate leaching step further results in the formation of K ₂ SO ₄ . In such instances, it is preferred that the Mn-lean liquor further includes K ₂ SO ₄ and residual Mn in a K:Mn molar ratio of from about 0.5: 1 to about 2: 1; and the process further includes: contacting the Mn-lean liquor with an oxidant to oxidise residual Mn in the Mn-lean liquor to heptavalent Mn while adding a neutralising agent to the Mn-lean liquor to adjust or maintain the pH of the Mn-lean liquor to a value of from pH 4 to pH 8; crystallising KMnOr in the Mn-lean liquor to form a slurry of KMnOr crystals; and separating the slurry into a KMn0 ₄ containing solids stream and a liquids stream.

[0029] In one form of the above embodiment, the oxidant is selected from the group consisting of: hydrogen peroxide, ozone, sodium persulphate, potassium persulphate, and Caro’s acid.

[0030] In one form of the above embodiment, the neutralising agent is selected from the group consisting of: sodium hydroxide and potassium hydroxide.

[0031] In one form of the above embodiment, the pH of the Mn-lean liquor is adjusted to a value of from pH 4 up to pH 7. It is preferred that the value is from pH 4.5. More preferably, the value is from pH 5. Most preferably, the value is from pH 5.5. Alternatively or additionally, it is preferred that the value is up to pH 6.8. Most preferably, the value is up to pH 6.5.

[0032] In one form of the above embodiment, the step of contacting the Mn-lean liquor with an oxidant is conducted at a temperature of from ambient to l00°C. Preferably, the temperature is from about 40°C. More preferably the temperature is from about 45°C. Most preferably, the temperature is from about 50°C. Additionally or alternatively, it is preferred that the temperature is up to about 90°C. More preferably, the temperature is up to about 80°C. Even more preferably, the temperature is up to about 70°C. Most preferably, the temperature is up to about 60°C.

[0033] In one form of the above embodiment, wherein the liquids stream includes residual K2SO4, the process further includes: converting the residual K2SO4 to KOH and H2SO4.

[0034] In one form of the above embodiment, the step of converting the K2SO4 includes using bipolar cell to split the liquids stream into a KOH containing fraction and an H2SO4 containing fraction. [0035] In a fourth aspect of the invention, there is provided a process for separating Mn from a Mn-containing leachate, the process including: contacting the Mn-containing leachate with an alcohol to form an Mn-precipitate and a Mn-lean liquor; and separating the Mn-precipitate and the Mn-lean liquor.

[0036] In an embodiment, the Mn-containing leachate includes Mn in the form of MnSCU. Preferably, the Mn-containing leachate is a MnSCU leachate, and the Mn-precipitate is a MnSCU precipitate.

[0037] In an embodiment, the Mn-containing leachate is formed by: leaching the Mn-containing feed with sulphuric acid and a reductant to reduce Mn in the solid feed and form a nSCU containing leachate and an Mn-lean solid residue; and separating the nSCU containing leachate and the Mn-lean solid residue.

[0038] In an embodiment, the reductant is a reducing sulphur species. Preferably, the reducing sulphur species is SCU.

[0039] In accordance with a broad aspect of the present invention, a process for the recovery of pure manganese products, with manganese sulphate as a precursor, and including the recovery of potassium and calcium as value-added by-products, is described.

BRIEF DESCRIPTION OF THE DRAWINGS

[0040] Figure 1: A process flow diagram for the removal of calcium from a Ca- and Mn- containing solid according to embodiments of the first aspect of the invention.

[0041] Figure 2: A process flow diagram for the recovery of manganese according to embodiments of the first or second aspects of the invention.

[0042] Figure 3: A photograph of MnSCU crystals recovered in Example 3.

DETAILED DESCRIPTION OF EMBODIMENTS OF THE INVENTION

[0043] The embodiments of the present invention shall be more clearly understood with reference to the following detailed description taken in conjunction with the accompanying drawings, Figure 1 and Figure 2.

[0044] The description which follows, and the embodiments described therein are provided by way of illustration of an example of particular embodiments of principles and aspects of the present invention. These examples are provided for the purposes of explanation and not of limitation, of those principles of the invention. In the description that follows, like parts and/or steps are marked throughout the specification and the drawing with the same respective reference numerals.

[0045] Referring to Figure 1 , there is shown a schematic representation of a leaching circuit. Many manganese-bearing ores also contain significant amounts of calcium and of potassium, and for the purposes of this description, such an ore has been selected.

[0046] The feed 10 is leached 11 with recycled hydrochloric acid 12. The purpose of this leach is to remove contained calcium without dissolving any manganese. Consequently, an acid with a corresponding soluble calcium salt is used; the leach conditions are mild with dilute acid, the strength and amount of acid being dependent on the amount of calcium present. The leaching temperature may be from ambient to l00°C, but preferably in the range of 50-60°C. The pH of the resultant solution is controlled to be in the range 4.0-6.5, preferably 5.0-5.5 in order to facilitate the subsequent precipitation of calcium. Leaching may be carried out in any suitable vessel, such as, but not limited to, a column or a stirred tank and co-current, or more preferably, counter-current, the latter ensuring the relatively high pH of the final liquor. [0047] The leach slurry 13 proceeds to solid-liquid separation, which may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter.

[0048] The filtrate 15 is sparged with recycled SO2 gas 17 in accordance with a process 16 described in a previous invention of one of the current applicants, Carl W. White, Jean Guimont, Denys Pinard and Serge Monette, Process and Apparatus for Treating Foundry Sludge to Recover Magnesium, US Patent 6,409,980, issued on June 25, 2002, and incorporated herein by reference. The calcium sulphite slurry 20 so-produced proceeds to solid-liquid separation 21 which may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter. The precipitation reaction is shown in Equation 1.

CaCk + SO2 + H2O CaS0 ₃ + 2HC1 (1)

[0049] The calcium sulphite solids 20 are calcined at a temperature of 500-l000°C, preferably 700-800°C to generate a pure, reactive form of lime 22. The SO2 17 generated is recycled to the precipitation reactor 16. The decomposition reaction is shown in Equation 2.

CaSOs CaO + SO2 (2)

[0050] The solids 23 from the first leach filter 14 undergo a reducing leach 24 with recycled sulphuric acid 26 and SO2 25, in accordance with minimising the amount of dithionates generated, familiar to those skilled in the art. The leach may be carried out in any suitable vessel such as, but not limited to, a column or a stirred tank reactor.

[0051] The leach slurry 27 proceeds to solid-liquid separation 28 which may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter. The solids 29 are sent to disposal, and the filtrate proceeds to manganese recovery as per Figure 2.

[0052] Turning now to Figure 2, there is shown a schematic for the recovery of various manganese products. The reducing leach solution 30 from Figure 1 is mixed 31 with recycled methanol (CH3OH) 32 in order to crystallise a very pure form of manganous sulphate 35. It has been discovered that methanol, rather than higher-chain alcohols, is very effective at selective crystallisation for manganese. For instance, if copper is present, it remains in solution, whereas if ethanol, for example, is used, then a mixed manganese-copper crystal is obtained. Such a method is more effective than, for instance, the traditional fractional crystallisation method.

[0053] The amount of methanol added is dependent upon the concentration of the manganese present, but a typical ratio is 1 : 1 methanol to leach solution. The crystal slurry proceeds to solid- liquid separation 34 which may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter.

[0054] The crystals 35 are pure manganese(II) sulphate of indeterminate waters of crystallisation, and denoted as MnSCU'xFFO. This is a product of its own accord, but may optionally be used as the precursor for producing other manganese chemicals 36 such EMM (electrolytic manganese metal), manganese oxide (MnO), manganese carbonate (MnC0 ₃ ), EMD or CMD (electrolytic or chemical manganese dioxide, Mn0 ₂ ).

[0055] The filtrate 37 is heated 38 to 50-l00°C, preferably 80°C, in order to distill off the methanol 32 for recycle and re-use.

[0056] The remaining aqueous liquor 39 may be treated for recovery of potassium if present. If not, it simply recycled to the reducing leach. If potassium is present, then the amount of methanol added at 31 is limited such as to leave sufficient manganese in solution to react with the potassium, being a molar ratio of 1: 1 K:Mn.

[0057] The liquor is first oxidised 40 with a suitable oxidant 41, such as, but not limited to, hydrogen peroxide in order to convert the manganese to its heptavalent (VII) state. Simultaneously, the solution is neutralised with recycled potassium hydroxide 42 to a pH value of between 4 and 8, preferably between 5 and 7, and more preferably pH 6.0.

[0058] The oxidised and neutralised liquor 43 is then crystallised 44 in any suitable crystalliser, which may be, for example, a vacuum, cooling or evaporative crystalliser. The crystal slurry proceeds to solid-liquid separation 45 which may be effected by any convenient means, such as, but not limited to, flocculation and thickening, filter press or vacuum belt filter. The solids are potassium permanganate 46. [0059] The liquid 47 is potassium sulphate, and may be treated for salt splitting 48 in a bipolar cell. This permits recycle of potassium hydroxide 42 to oxidation/neutralisation 40 and sulphuric acid 26 to the reducing leach of Figure 1.

[0060] The process will now be illustrated by way of examples. These examples are provided for the purposes of illustration of certain embodiments of the invention, and are not intended as limitations thereof.

[0061] Example 1

A 100 g sample of manganese magnetic concentrate, analysing 52.8% Mn, 0.74% Ca and 0.16% Cu was leached in sulphuric acid at 90°C for 4 hours. Hydrogen peroxide was added, but as a reductant rather than oxidant. The final ORP of the slurry was 68 mV at room temperature. Under these conditions, manganese extraction was 90%, but copper extraction was only 30%. This test, however, indicates that manganese is readily leachable under reducing conditions.

[0062] Example 2

A series of leach tests similar to that of Example 1 were carried out, except that the slurry was then neutralised with lime to raise the pH to 4.0 in two tests, and 4.5 in the other 2, after which the slurry was filtered resulting in similar manganese and copper extractions. The filtrate pH was raised 6.5 with caustic addition, with hydrogen peroxide, acting as an oxidant this time, being added simultaneously. The temperature was then raised to 90°C, resulting in the formation of a brown precipitate. This was repeated four times, twice with the pH 4 filtrate (samples 1 and 2) and twice with the pH 4.5 filtrate (samples 3 and 4), with the final solids analysis as in the Table below.

XRD analyses of the precipitates showed mixed manganese oxides, as would be expected. The data show that a high-quality mixed manganese oxide can be recovered in this manner.

[0063] Example 3

The leach test of Example 1 was repeated, and the slurry neutralised to pH 4.5 with 20% lime slurry as in Example 2. One litre of solution was taken, analysing 24 g/L Mn, to which 1.4 L of methanol was added at room temperature. Coarse, slightly pink, translucent crystals were obtained, characteristic of MnS0 _4* H ₂ 0, which after washing in methanol and drying yielded 33 g. Based on these being the monohydrate as suggested by their colour, this represents a recovery of 27% of the manganese, but more importantly indicates that a residual solubility of manganese in solution of about 17 g/L is readily attainable. Thus, much higher recoveries would be attained form more concentrated solutions than that used in this example. Figure 3 is a photograph showing the produced MnSCE crystals.