The present invention relates to a process for the production of electrolytic manganese dioxide. More particularly, the process for the production of 5 electrolytic manganese dioxide of the present invention may additionally produce a soil conditioner or improver.

Background Art

The traditional process route for the production of electrolytic manganese dioxide (hereinafter "EMD") involves firstly the pyrometallurgical reduction of a high

10 manganese dioxide content ore, typically with a manganese dioxide content of over 40% by weight, to produce a mono-oxide of manganese, MnO. The MnO is then leached in sulphuric acid to produce a liquor containing MnS0 ₄ . Impurities are then removed from this liquor, EMD is deposited on titanium anodes by way of an electrochemical process, for example electrowinning, and that deposited EMD

15 is then scrapped, neutralised, milled and dried.

The removal of impurities from the MnS0 ₄ liquor traditionally produces a waste product containing significant levels of unreacted MnO, and also both of calcium sulphate and ferric hydroxide, which are ultimately simply allowed to remain as tails of this process. The loss of MnO to the process obviously impacts on the 20 efficiency of the process.

Simultaneously, there is an estimated current worldwide demand for trace elements fertilisers or soil conditioners/improvers and a growth in that demand ^~~~ exceeding 7% per ^' annum The

and environmental factors affecting the availability of micronutrients to the plants 25 in developed countries. The major factors include pH, organic matter content, nutrient interaction, soil water content, temperature and light. For example, because of high and constant use of phosphate fertilisers in the developed countries, the soil pH has increased to a level where availability of B, Cu, Fe, Mn and Zn has decreased dramatically.

The International Fertiliser Industry Association (IFIA) estimates the worldwide consumption of macron utrients fertiliser (NPK) in 2008 at 170 million tonnes (Short Term Fertilizer Outlook 2009-10, IFIA). Though no exact quantities of micronutrients fertiliser consumed are provided by any organisation, IFIA estimates that the consumption of micronutrient fertiliser is around 2 to 2.5% of macronutrient fertilizer consumption. The Applicant estimates that the world consumption of micronutrients fertilizer is at 2% of macronutrient fertiliser in the year 2008, thus giving a total consumption of micronutrients fertiliser in 2008 at around 3.4 million tonnes per annum. Australia is reported to import various compounds containing these trace elements in excess of 200,000 MT/ annum. Usually such products are imported as a compound of individual element such as manganese sulphate, zinc sulphate, sodium molybdate, borax and the like. There are very few products available that contain a mixture of several trace elements suitable for Australian and international soils.

The process of the present invention has as one object thereof to overcome the abovementioned problems associated with the prior art, or to at least provide a useful alternative thereto.

The preceding discussion of the background art is intended to facilitate an understanding of the present invention only. The discussion is not an acknowledgement or admission that any of the material referred to is or was part of the common general knowledge as at the priority date of the application. Throughout the specification and claims, unless the context requires otherwise, the word "comprise" or variations such as "comprises" or "comprising", will be understood to imply the inclusion of a stated integer or group of integers but not the exclusion of any other integer or group of integers. ^" Πιβ terms soil conditioner, soil improver and fertiliser are to be understood as designating essentially the same product unless the context demands otherwise.

Disclosure of the Invention

In accordance with the present invention there is provided a process for the production of electrolytic manganese dioxide, the process characterised by the method steps of:

(i) Reducing a manganese dioxide containing ore to produce a manganese oxide product;

(ii) Leaching the manganese oxide product with sulphuric acid producing a manganese sulphate containing liquor;

(iii) Passing the liquor from step (ii) to a jarositing step producing a precipitate;

(iv) Passing the liquor from step (iii) to a goethiting step producing a precipitate; (v) Passing the liquor from step (iv) to a sulphiding step and filtering the resulting liquor to remove heavy metal impurities as heavy metal sulphides;

(vi) Passing the remaining liquor from step (v) to an electrowinning step in which electrolytic manganese dioxide is deposited and a spent liquor produced; and

(vii) Recycling spent liquor from step (vi) directly or indirectly to the leach of step (ii) to leach manganese oxide thereby contributing favourably to the efficiency of the process. Preferably, the precipitates of steps (iii) and (iv) are used to produce a soil conditioner.

Still preferably, the precipitate from step (iii) is passed to a dense media separation step in which silica is removed before combination with the precipitate of step (iv) in the production of the soil conditioner.

The combination of the precipitates from steps (iii) and (iv) are preferably combined in a filter step that produces a generally solid- product that is in turn used in the production of the soil conditioner. The remaining liquid from the filter step is preferably passed directly or indirectly to the leach of step (ii) to the manganese oxide slurry.

Preferably, the solid product of the filtration of the precipitates of steps (iii) and (iv) has added thereto one or more trace elements. The solid product is preferably also dried and granulated in the production of the soil conditioner.

The added trace elements may be provided in the form of one or more micronutrient compounds. The micronutrient compounds preferably include one or more of zinc sulphate, copper sulphate, borax, and sodium molybdate.

The process of the present invention further preferably comprises the following additional method steps:

(viii) Scrapping of the anodes from the electrowinning step (vi) to remove the deposited electrolytic manganese dioxide; and

(ix) Neutralising and milling the scrapped electrolytic manganese dioxide.

In accordance with the present invention there is further provided a soil conditioner characterised in that it comprises precipitated material from the jarositing and goethiting of a manganese sulphate liquor. Preferably, the manganese sulphate liquor is produced in the leaching of a manganese oxide material with sulphuric acid.

Still preferably, silica is removed from the precipitated material from jarositing.

One or more trace elements may be combined with the precipitated material from jarositing and goethiting. The added trace elements may be provided in the form of one or more micronutrient compounds. The micronutrient compounds preferably include one or more of zinc sulphate, copper sulphate _t borax, and sodium molybdate.

In one form of the invention the soil conditioner is provided in dried and granulated form.

Brief Description of the Drawings

The process for the production of electrolytic manganese dioxide of the present invention will now be described, by way of example only, with reference to one embodiment thereof and the accompanying drawing, in which:- Figure 1 is a schematic flow sheet of a process for the production of electrolytic manganese dioxide in accordance with the present invention, the flow sheet also showing how precipitated material from jarositing and goethiting steps may be used to produce a soil conditioner.

Best Mode(s) for Carrying Out the Invention In Figure 1 there is shown a process 10 for the production of electrolytic manganese dioxide ("EMD").

The reduction of Mn0 ₂ ore, most commonly pyrolusite, to MnO is an essential requirement for the hydrometallurgical processing of Mn0 ₂ ore in order for it to be leachable with sulphuric acid. That is, manganese is reduced from its 4+ valence state to its 2+ valence state to facilitate its dissolution in subsequent hydrometallurgical processing. Similarly, whatever iron oxides are also present will be reduced, whereby iron will be present in predominantly the 2+ valence state. Accordingly, a Mn02 containing concentrate 12 is mixed with coke 14 (as a source of carbon) in a ratio ranging from 1 :0.1 to 1 :0.5 and heated at about 1000°C by compressed natural gas ("CNG") in a reduction furnace 16 to reduce the dioxide ore to mono-oxide of manganese.

Calcination takes place at about 1000°C in the reduction furnace 16 with a residence time of between about 15 minutes to one hour (at temperature), during which the Mn0 ₂ is converted to MnO. The converted product of the reduction furnace is passed to a calcine cooler 18.

Off-gases from the reduction furnace 16 are cleaned in an electrostatic precipitator (not shown) before discharging the off-gases to the atmosphere via a stack (not shown). Recovered solids from the off-gases are conveyed to the calcine cooler 18.

The cooled product from the calcine cooler 18 is then milled in a closed circuit wet grinding mill 20 to reduce the material to a Peo of 75 μιτι and the mill product is then pneumatically conveyed to a storage silo (not shown). This fine size range is preferred to ensure high Mn recoveries in the subsequent process steps. The reactions taking place are as follows:

2Mn0 ₂ + C = 2MnO + C0 ₂

2Fe ₂ 0 ₃ + C = 4FeO + C0 ₂

Leaching takes place in continuous leach tanks 22 which receive solids from the wet grinding mill 20, spent sulphuric acid liquor from a spent sulphuric acid liquor tank 24, optionally sulphuric acid from sulphuric acid overhead tanks 26 and make up water from process water pumps (not shown). The pH of the leach is about 1 to 1 .5, for example about 1. A pulp density of 40% solids (w/w) is used for the leach to maximise the Mn concentration and to ensure that the solubility limit of Mn sulphate during leaching is not exceeded.

The leach tanks 22 operate at ambient pressure and a temperature of about 90°C with a nominal tank residence time of about 1 hour. The leach provides manganese and iron recoveries of greater than 90%.

The Mn sulphate liquor and residue slurry exiting the. leach tanks 22 are pumped to a jarositing step, comprising jarositing tanks 28, for impurity removal.

The reactions taking place are understood to be as follows: R1 : MnO ^■■ + H ₂ S0 ₄ = MnS0 ₄ + H ₂ 0 100%

R2: K ₂ 0 + H ₂ S0 ₄ = K ₂ S0 ₄ + 4H ₂ 0 100%

R3: FeO + H ₂ S0 ₄ = FeS0 ₄ + H ₂ 0 100%

R4: 2AIOOH + 3H ₂ S0 ₄ = AI ₂ (S0 ₄ ) ₃ + 4H ₂ 0 30%

R5: MgO + H ₂ S0 ₄ = MgS0 ₄ + H ₂ 0 85% R6: H ₂ 0 ₍ g ₎ = H ₂ 0 ₍ i ₎ 100%

R7: H ₂ 0 ₍ i ₎ = H ₂ 0 ₍ g ₎

The first precipitation stage, the jarositing step, removes potassium by increasing the pH of the manganese sulphate liquor to about 1.8 to 2 by adding limestone slurry. The exact reaction taking place is not known but it is believed that potassium forms an insoluble complex with iron. The aim of jarosite precipitation is the removal of potassium, preferably to a level less than 5 mg/L so as to prevent the formation of Cryptomelane which would be detrimental to the ultimate EMD product. Jarosite precipitation is conducted with an Fe/K ratio of about 12:1 , although higher ratios of 12:1 may be utilised to ensure appropriate levels of potassium removal.

The product of the jarositing step is passed to a thickening step 30 from which the , thickener overflow is passed to a second precipitation or goethiting step 32.

The second precipitation stage 32 removes Al, Fe and various other impurities. This is achieved by increasing the pH of the n sulphate liquor to between 4 to 5 by adding hydrated lime slurry. The following reactions are understood to take place: R1 : Fe ₂ (S04)3 + 3Ca(OH) ₂ = 2Fe(OH) ₃ + 3CaS0 ₄ 100%

R2: AI ₂ (S0 ₄ ) ₃ + 3Ca(OH) ₂ = 2AI(OH) ₃ + 3CaS0 ₄ 100%

R3: H ₂ S0 ₄ +Ca(OH) ₂ = CaS0 ₄ + 2H ₂ O ₎ 100%

R5: H ₂ 0(i) = H ₂ 0 _(g) 100% A third precipitation stage or sulphiding step 34 receives overflow from a thickening step 36 that receives the product of the goethiting step 32. The sulphiding step 34 removes heavy metal impurities by sulfiding the manganese sulphate liquor. The manganese sulphate liquor is treated with barium sulfide to precipitate heavy metal impurities as sulphides. The resulting liquor is passed to a filter, for example a belt filter 38, to obtain a manganese sulfate liquor to process electrochemically to deposit EMD on titanium anodes.

The precipitation circuit, incorporating the jarositing step 28, the goethiting step 32 and the sulphiding step 34, is designed to remove potassium, iron, aluminium and calcium impurities, and heavy metal impurities, from the manganese sulphate solution and provide a clear, high purity filtrate for EMD electrowinning.

EMD deposition from the manganese sulphate liquor takes place in an electrowinning step 40 on titanium anodes (using copper cathodes) in cell houses. The manganese sulphate liquor containing about 60 to 70 g/L manganese is pumped to an electrochemical cell house with a sufficient amount of sulphuric acid to maintain pH at about 1. The electro-deposition of EMD takes place at about 90°C with current density of about 60 Amp/m ² . The following reaction is understood to take place at the anode: MnS0 ₄ + 2H ₂ 0 = Mn0 ₂ + H ₂ S0 ₄ + H ₂

The EMD deposited on titanium anodes is scrapped off from the anodes in a scrapping step 42. The EMD is neutralised using soda ash solution in a first neutralisation step 44, then wet ball milled 46, then passed to a second neutralisation step 48 and again wet ball milled 50, to the desired customer specifications, dried 52 and packaged 54.

The spent liquor from the electrowinning step 40 is passed to the spent liqubr tank 24 before being fed back to the leach step 22, either directly or via the wet grinding mill 20, to leach manganese oxide in the leach step 22. The spent liquor contains significant amounts of sulphuric acid and some manganese sulphate. Generally, this recycled sulphuric acid is sufficient to achieve the necessary leaching of manganese oxide in the leach step 22, thereby contributing to the efficiency of the process 10 of the present invention.

The thickener underflow from the thickener 30 is passed to a dense media separation ("DMS") step 56 in which silica is separated from the other precipitated solids from the jarositing step 28. The float from the DMS step 56 is filtered in a filter step 58, using for example a belt filter. The thickener underflow from the goethiting step 32 and thickener 36 is also passed to this filter step 58. The liquid product of the filter step 58 is passed to the leach step 22 either directly or via the wet grinding mill 20, again so as to leach manganese oxide.

The solids from the filter step 58 are passed to a drier 60 and in turn to a granulator 62. Additives in the form of trace elements are added to the product from the drier 60 as it passes to the granulator 62. These trace elements may be provided in the form of one or more micronutrient compounds. The micronutrient compounds include, but are not limited to, zinc sulphate, copper sulphate, borax, and sodium molybdate.

It is envisaged that while producing high alkaline grade EMD, the process of the present invention can also produce a soil conditioner that contains Mn, Fe, Ga, and S in sufficient quantities and in a form that is bioleachable, or available, for the plants. That is, the form in which the elements are present is such that they are available to plants to be taken up. As noted above, the basic product of the jarositing step 28 and the goethiting step 32 may be supplemented by blending with some other important trace elements, additives, not otherwise present, but that are considered necessary for plant growth. Whilst there is scope to vary the specific composition of the soil conditioner, Table 1 below contains one preferred example thereof:

Table 1

Elements (%)

Mn >4.00

Fe >4.00

Ca >6.00

S >9.00

Zn >4.00

Mo >0.05

Cu >4,00 .

B >2.00 Mn is present as mono-oxide of Mn (MnO), Fe as ferric oxide, Ca as calcium sulphate, Zn as zinc sulphate, Cu as copper sulphate, Mo as sodium molybdite, and B as borax. These elements are considered to be the essential trace elements required for the proper growth of crops. For example, Mn is involved in regulation of enzymes and growth hormones. Mn also assists in photosynthesis and respiration. Fe is involved in respiration and chlorophyll synthesis. Ca is involved in formation of cell walls, and root and leaf development. Ca also participates in translocation of sugars. S is involved in formation of nodules and. chlorophyll synthesis, structural components of amino acids and enzymes. Zn is involved in production of growth hormones and chlorophyll. It also assists in respiration and carbohydrate synthesis. Mo is involved in enzymatic reduction of nitrates to ammonia. Mo also assists in conversion of inorganic phosphates to organic form. Cu is involved in photosynthesis and respiratory system. Cu also assists in chlorophyll synthesis and used as a reaction catalyst. Boron is involved in formation of cell walls, terminal buds and pollen tubes. B also participates in regulation of starch production and translocation of sugars and starches.

Australian, American and European soils are largely deficient in these trace elements. There is significant consumption of these trace elements in these countries. For example, it is estimated that these countries use around 3.4 million tonnes of these micronutrients every year.

Further, there is an estimated current worldwide demand of these trace elements at 3.4 million tonnes per annum with a growth exceeding 7% per annum. The growth is high because of soil and environmental factors affecting the availability of micronutrients to the plants in the developed countries. The major factors include pH, organic matter content, nutrient interaction, soil water content, temperature and light. For example, because of high and constant use of phosphate fertilisers in the developed countries, the soil ]5H~has increased to a level where availability of B, Cu, Fe, Mn and Zn has decreased dramatically. The soil conditioner of the present invention, containing S as sulphates of Ca, Zn, and Gu, will act as a soil conditioner and remedy the soil pH. This will enable the plants to take up these micronutrients.

The process for the production of electrolytic manganese dioxide of the present invention may be better understood with reference to the following non-limiting examples.

Example 1 : Overall Process for EMD Production

Concentrate feed composition, being manganese ore from the Groote Eyiandt Mining Company Pty Ltd (GEMCO), a BHP Billiton subsidiary, project:

STEP 1 : PYROMETALLURGY

1.10 Reduction of Ore (Production of EMD) Experimental i) Prepare 3 x 1 kg sample of the provided particle size <(-6 mm) and

Coal to Natural Gas Stoichiometric ratio at 0.5 to 1.0. ii) Run these three samples independently at 1000°C temperature and residence time of 30 minutes. iii) Analyse all the three samples independently by XRD (or any other analytical tool) to determine reduction of Mn02 to MnO. iv) Use these three samples independently for hydrometallurgical processing to produce manganese sulphate.

STEP 2: H YD ROM ETALLU RG Y

2.10 Milling of Reduced Ore (Number of Samples: 3) Experimental i) Wet ball milling of each 1 kg sample after reduction particle size to Peo 75pm and 40% w/w solkTslurry. Expected SG 1.41 , slurry weight 2.5 kg, slurry volume 1.77 L.

2.20 Washing of Milled Reduced Ore (Number of Samples: 3) Experimental i) Use 40% w/w pulp slurry from the wet ball mill for each sample (2.5 kg) and stir for 15 minutes. ii) Add 15% water of slurry weight (375 g) and filter at 5 kPa vacuum. iii) Repeat ii) above. iv) Repeat iii) above. v) Assay filtrate and residue after step iv) for potassium. vi) Report potassium removal by washing. vii) Dry the samples at 105°C overnight to remove moisture. Report the moisture inherent in the wash residue. 2.30 Leaching of Milled Reduced Ore and Goethiting after Step 2.20 (Number of Samples: 3)

Experimental i) Leach each sample separately-with sulphuric acid (98%) at 95°C for- a residence time of 3 hours as per the following conditions. a) Make a slurry 40% w w solids. Water approximately 1.5 L to be added. Expected SG of slurry 1.41. Measure the SG after making 40% solids w/w slurry. b) Add 3 L of water. c) Add 1.29 kg sulphuric acid (98%) slowly by maintaining temperature at 95°C for a residence time of 3 hours. Make sure that pH does not exceed 1.5 at any time. d) After step c), cool the leach solution. e) Allow to settle Pregnant Leach Solution (PLS), measure pH, and assay PLS for all the elements using XRF (or any other suitable analytical tool). We expect Mn ²⁺ concentration in the PLS at approximately 80 to 90 g/L. ii) Jarosite and goethite the slurry to remove K, Fe, Al etc. a) Stir the PLS solution with solids and get slurry. Heat it to 95°C. b) Add hydrated lime 30% solution to the slurry by constant stirring and maintaining temperature at 95°C for 1 hour until pH reaches 1.8 to 2. Measure the quantity of hydrated lime to achieve pH 1.8 to 2 in the slurry. c) Add H ₂ 0 ₂ to oxidise ferrous to ferric until SHE 800 mV.

Measure the quantity of H2O2 added to arrive at SHE 800 mV. d) Repeat step b) until pH reaches about 4 to 5. e) Add Magnofloc™ flocculant from NALCO™ (or other suitable flocculant) at suitable rate (preferably at approximately 0.060 g/L of slurry) and allow slurry to settle. Report amount of flocculant required to efficiently settle the slurry and residence time. f) Decant the PLS, and assay for all the elements using

XRF (or any other suitable analytical tool). g) Filter the residue three times by using approximately 15% water w/w of the residue (approximately 1L water) at each time. Send residue to dry and add additional trace element compounds and granulate. (h) Expected volume of slurry 5.49 L, SG approximately

1.24. iii) Sulfiding of PLS. a) Calculate stoichiometrically the amount of BaS required to precipitate heavy metal impurities from the assay results of PLS as above. b) Heat PLS at 95°C. c) Add above calculated BaS to PLS. d) Cool the PLS and filter to remove sulphides. e) Assay PLS for all the elements using XRF (or any other suitable analytical technique). Heavy metals should be within the expected limits. ^{* ~"} iv) Storing of PLS. a) Store the PLS for electrometallurgy treatment. Expected Mn ²⁺ concentration is 80g/L.

STEP 3 ELECTROMETALLURGY i) Adjust Mn ²⁺ concentration of PLS to approximately 65 g/L (±5) Mn. ii) Add 45 g/L sulphuric acid 98% to the above. iii) Heat the above solution to 95°C. Referred to here as electrolyte. iv) Subject the above electrolyte for electrodeposition of Mn02 on Titanium grade 2 anodes and Cu cathodes for a period of 10 to 14 days (subject to the laboratory condition of the lab). Temperature of cell house 95°C (±5°C) Current 60 Amp/m ² , Voltage average 3.2 to 3.6. v) Make sure that the concentration of Mn ²⁺ in the electrolyte in the cell house does not fall below 45 g/L. v) After the completion of deposition cycle, take out the Ti electrode and wash with deionised water. vi) Scrap the EMD from the Ti electrodes. vii) Wet mill first to D ₅ o 35pm and neutralise using 10% caustic soda solution. viii) Wash 3 times with 15% w/w water of EMD slurry to clear slurry with any excess caustic soda and/ or sodium sulphate. viii) Wet mill EMD now to D ₅₀ 2pm. ix) Again neutralise wet slurry of EMD by caustic soda 10% solution. Expected pH 6.5. ^' ' x) Wash 3 times with 15% w/w water of EMD slurry to clear slurry with any excess caustic soda and/ or sodium sulphate. xi) Dry and weigh the EMD powder. xii) Determine electrochemical activity of EMD powder for alkaline batteries.

Example II

Tests were conducted to investigate leach and precipitation conditions as follows: A reduced ore of the composition shown in Table 2 below was used as the basis of several leach and precipitation tests:

Table 2: Chemical composition of the reduced ore (%, w/w)

Initial Leach Tests The ore was mixed with deionised water to make up a slurry of 40% solids (w/w) prior to feeding into the reactors. The density of the 40% slurry was measured at 1.4 kg/L. A bulk synthetic solution representing the spent electrolyte was prepared with a concentration of 70 g/L sulphuric acid and 38 g/L manganese using AR grade sulphuric acid (98%) and manganese sulphate salt. The tests were conducted in 29 L reactors fitted with baffles for mixing and steam coils for heating. The reactors were filled with 21.35 L of synthetic solution and heated to 90°C before feeding the ore slurry. The solids concentration in the reactor was 8% (w/w) after adding the ore slurry, which contained 3.15 L of deionised water. The ore slurry was generally added within the first 30 minutes of the leach. The pH was first controlled at 3 by adding sulphuric acid solution.

The pH in the first leach test (L1 ) was adjusted by adding a 10 M (980 g/L) sulphuric acid solution, which was prepared by diluting concentrated sulphuric acid (98%) with the synthetic spent electrolyte. Dilution of concentrated acid with the synthetic electrolyte resulted in an increased manganese concentration in the PLS. As a result, the bulk synthetic electrolyte prepared was diluted with deionised water and used as a lixiviant for L2 and L3. The acid and manganese concentrations after dilution were 44 g/L and 24 g/L, respectively. A 9M sulphuric acid solution was used for pH adjustment in L2. This solution was prepared by diluting concentrated sulphuric acid (98%) with deionised water decanted from the ore slurry and a small volume of the diluted synthetic solution. The pH during L3 was adjusted by addition of a 9 M sulphuric acid solution, where concentrated sulphuric acid (98%) was diluted with deionised water decanted from the prepared ore slurry. Bulk leach tests

The filtrate from BLI was 46 L and it was reduced to 31 L after jarosite/goethite precipitation due to evaporation. This solution was less than the minimum of 50 L required to run electrowinning for 14 days. It was therefore, necessary to conduct another bulk !each test (BL2) to produce sufficient solution. The filtrate from BL2 68 L and it was reduced to 53 L after jarosite/goethite precipitation. The solution from BL1 and BL2 were combined after jarosite/goethite precipitation to conduct sulphide precipitation. After sulphide precipitation, some elements did not meet the specification required for the electrowinning solution. A third bulk leach, BL3, was conducted to produce a fresh PLS that could be processed for potassium removal by jarosite precipitation, as previous results had shown that potassium removal was efficient when jarosite precipitation was conducted on a fresh PLS directly after leaching. The ore used in BL3 was milled separately from the ore previously used. The three bulk leach tests (BL1 , BL2 and BL3) were conducted in a 150 L reactor, fitted with baffles and steam coils for heating. The only variable in the bulk leach tests was the % solids (w/w), which were 6%, 4% and 3.5% for BL1 , BL2 and BL3, respectively. The same procedure was followed for all tests. The mass of ore leached varied slightly but was approximately 3 kg for each test. A fresh bulk synthetic electrolyte was prepared as a lixiviant with a composition resembling the spent electrolyte from the electrowinning circuit. The manganese and sulphuric acid concentrations were 35 and 30 g/L, respectively. The ore was fed as a 40% solids (w/w) slurry, and pH was controlled at 1.5 by adding concentrated sulphuric acid (98%). Concentrated sulphuric acid was used to avoid the slow pH adjustment observed when using diluted acid. The use of diluted acid would also result in larger volumes of acid solution required for pH adjustment at larger test scale which would lead to dilution especially at the lower % solids. After each leach test, the slurry was filtered and the PLS was stored for subsequent jarosite/goethite precipitation. Results for all leach tests conducted are shown in Table 3 below. The first three leach tests, i.e L1 , L2 and L2 were conducted at pH 3 and all other tests, i.e BL1 , BL2 and BL3 were conducted at pH 1.5. All tests were conducted at 90°C and pH measurements were recorded at temperature. During L1 , some slurry was lost due to the violent reaction when adding acid to adjust the pH while introducing ore into the reactor. Thereafter, the ore and acid were added at a slower rate to avoid spillage of slurry. Samples from the individual filtrate solutions from L1 , L2 and L3 were not analysed but a sample of the combined filtrate solutions was submitted ^™ for analysis. Analysis of the filtrate solutions from the individual tests was conducted to compare kinetics for the three leach tests. The recovery for the individual tests was not calculated. The potassium, manganese and iron concentrations in the combined filtrate sample were 709 rng/L, 84 g/L and 5.2 g/L, respectively.

The Manganese and iron recoveries for the bulk leach tests, BL1 , BL2 and BL3 were > 90% and the highest recoveries for manganese and iron were 99% and 5 97%, respectively.

Table 3: Leach results

The average acid consumption calculated for L2, L3, BL1 and BL2 was approximately 1296 kg H2S04/ton ore. The low acid consumption for L1 could be 10 attributed to the slurry spillage when feeding ore and acid into the reactor. The reason for the unusually high acid consumption recorded for BL3 is unclear at this stage. However, the ore used for this test was from a separate reduced batch of ore that was milled separately from the first batch of ore used in the first 5 leach tests.

15 The results above favour conducting the leach at a lower pH than that originally _ . . _^ anticipated, at least as.low as pH -1.5. The Applicants believe that conducting the — leach at a pH of about 1 is likely to produce good results. Jarosite/Goethite precipitation

Jarosite/geothite precipitation tests were carried out by heating up the PLS to 90°C and maintaining the pH at 1.5 for jarosite precipitation unless stated otherwise. The aim of jarosite precipitation was to remove potassium in the PLS down to < 5mg/L to prevent formation of Cryptomelane, which is detrimental to the electrochemical activity of the EMD product. Jarosite precipitation was conducted for two to three hours and thereafter, the pH was increased to 4.5 to 5 for goethite precipitation for iron removal. Ferric sulphate was added to the PLS prior to jarosite precipitation to obtain an Fe/K ratio of 12/1. The only exception to ferric addition was the initial jarosite precipitation test (J1), where no ferric was added in an attempt to determine whether additional iron was required to remove potassium to < 5 mg/L by jarosite precipitation. Hydrated lime, and sometimes lime, was added as slurry to adjust the pH. An initial attempt was made during J1 to oxidise the ferrous using oxygen and to increase the redox potential to 550 mV. The maximum redox potential achieved using oxygen was 437 mV after 5 hours. EMD powder was added in an attempt to increase the redox potential, ~ 24 g of EMD was added, but the redox potential did not increase. Hydrogen peroxide was then added resulting in redox potential > 550 mV within minutes. Consequently, redox potential during all other jarosite precipitation tests was maintained at > 550 mV (Ag/AgCI) by addition of hydrogen peroxide. Some tests did not require addition of hydrogen peroxide due to the high ferric concentration after spiking the solution with ferric sulphate (anhydrous).

The initial jarosite precipitation test (J1) was conducted using the combined PLS from the first three leach tests (L1 , L2 and L3). The filtrate from J1 was spiked with ferric (as anhydrous ferric sulphate salt) to remove the remaining potassium; however, the residual potassium did not meet the specification. Potassium removal from this PLS was repeated ^" 5 times by spiking the filtrate from the previous jarosite/goethite precipitation test with ferric and conducting jarosite precipitation followed by goethite precipitation. The lowest potassium level attained was 15 mg/L. The filtrate solutions from BL1 , BL2 and BL3 were used separately to conduct three bulk precipitation; BP1, BP2 and BP3, respectively. BP1 and BP2 resulted in efficient potassium removal; however, BP3 was unsuccessful even after repeating jarosite precipitation on the filtrate twice. The filtrate solution from BP3 was combined with the sulphide precipitation (SP1) filtrate, which had a potassium concentration of 5 mg/L. This was done to obtain sufficient solution for electrowinning. The mixed solution was spiked with ferric sulphate and jarosite/goethite precipitation was conducted again in an attempt to lower the potassium to the required minimum; potassium removal to the required level was unsuccessful.

v

Jarosite precipitation was conducted first for a minimum of two hours at pH 1.8 (Tests J1, J2 and J3) and pH 1.5 (J4, J5, J6, BP1 , BP2 and BP3). Thereafter, the pH was increased by addition of lime slurry to pH 4.5 to 5 to conduct goethite precipitation for a minimum of 2 hours. Removal of potassium by jarosite precipitation proved to be inconsistent as indicated in the results in Tables 4 and 5 below. Potassium removal without ferric addition (J1) resulted in only 59% potassium removal at an Fe/K ratio of 7 and pH 1.8. Ferric sulphate (anhydrous) was added to the filtrate obtained from J1 to obtain an Fe/K ratio of 15at pH 1.8, this resulted in the removal of 80% of the residual potassium. The second repeat test (J3) was conducted at an even higher Fe/K ratio of 23 at pH 1.8, but no potassium was removed. The filtrate from J3 was adjusted to pH 1.5 in an attempt to improve potassium removal at an Fe/K ratio of 24, this resulted in the removal of 81% of residual potassium and a final potassium concentration. of 15 mg/L. Further processing of the solution by addition of excess ferric at pH 1.5 did not decrease the potassium content of the solution.

Table 4: Jarosite precipitation results for the initial leach solutions (90°C)

Initial K (mg/L) 660 268 80 82.5 15.3 17.5

Final K (mg/L) 268 80 82.5 15.3 17.5 31.25

% K removal · 59 70 81

Results for the removal of potassium from the bulk precipitation tests are shown in Table 5. Jarosite precipitation for the bulk precipitation tests was conducted at pH 1.5 and pH was measured at temperature. Potassium removal was efficient for BP1 and BP2; 99% of the potassium was removed from solution for both tests and the residual potassium concentration was within specification. Potassium removal from the BL3 filtrate was inefficient and only resulted in 19% potassium removal at an Fe/K ratio of 18. Jarosite precipitation was repeated twice on this solution and the minimum potassium level obtained was 74 mg/L even after increasing the Fe/K ratio to 37.

Table 5: Jarosite precipitation results for the bulk leach solutions (90°C, pH

The filtrate from SP1 , which was a sulphide precipitation test conducted using combined filtrate from BP1 and BP2, was mixed with the filtrate from BP3c to obtain sufficient solution to run the electrowinning test. This mixed solution was spiked with ferric sulphate and jarosite precipitation (BPM) was conducted in an - - attempt to further reduce the potassium level, this was followed by goethite precipitation for iron removal. Results indicated that no additional potassium was removed from solution; the test only resulted in the dilution of the potassium in the filtrate from BP3c by the filtrate from SP1. The increase in potassium concentration in BPM filtrate will be confirmed by additional analysis.

Whilst the above example was conducted at a lower pH, the Applicants anticipate conducting the jarositing step at a pH of about 1.8 to 2. Sulphide precipitation

The filtrate solution from BL1 and BL2 were mixed to make up approximately 83 L of solution for sulphide precipitation test (SP1). The client specified stoichiometric addition of BaS (99.7%, w/w) with an excess of 5% with respect to base metal impurities. However, stoichiometric amounts of BaS plus 12% excess was added twice in powder form to ensure complete removal of base metal impurities. The test was conducted at 90°C for 2 hours with no pH control. The filtrate was stored and the solids were processed; no moisture sample was taken in this case due to the low volume of solids produced. The solids were washed with deionised water by displacing the entrained solution with 1 L of deionised water three times, the filter cake was dried and the moisture content of the solids was calculated.

Solutions with the desired potassium concentration, i.e BP1 (31 L) and BP2 (52 L), were combined and sulphide precipitation was conducted (SP1). The test was conducted at 90°C and pH was not controlled, pH was measured at temperature. The initial pH of the solution was 4.91 and it increased slightly to 5.01 after addition of BaS. Stoichiometric mass of BaS plus -12% excess was added twice, first at 30 minutes and then at 1 hour. The stoichiometric mass was based on the concentration of the following elements: Ni, Zn, Co, Cu, Cd and Pb. Analytical results of the feed solutions (BP1 and BP2) and the filtrate from the sulphide precipitation test (SP1) were compared to the specification required for the - electrowinning solution. Only Co.-Cu, and Cd met the specification;- -The results indicated an increase in Ni, which is unusual. Table 6: Sulphide precipitation results

Analysis

After each acid leach and precipitation test, the filtrate sample was analysed. Solid samples will be analysed to complete the mass balance. The filtrate samples were submitted at the end of each test to determine the leach efficiency and extent of potassium removal for each test prior to proceeding with subsequent test work. The following elements were analysed by Inductively Coupled Plasma Optical Spectroscopy (ICP-OES) at a detection limit of 2 mg/L: Mg, Al, Si, Ca, Ti, V, Cr, Mn, Fe, Co, Ni, Cu, Zn, Mo, Pb, S, As, Li. Potassium in the filtrate samples was analysed by Atomic Absorption Spectroscopy (AAS). Base metals below 2 mg/L were analysed using Inductively Coupled Plasma Mass Spectroscopy (ICP-MS) to determine the BaS stoichiometric requirement. The detection limit for ICP-MS was 0,01 mg/L.

Kinetic samples from the first three leach tests (L1 , L2 and L3) were analysed for manganese and iron to determine leach kinetics. The synthetic electrolyte solutions were also analysed for manganese and sulphuric acid by titration. The concentration of ferric after spiking was checked by titration.

It is envisaged that the soil conditioner of the present invention will find ready acceptance from farmers and others working with fertiliser products. Amongst the expected advantages are:

(a) Most of the required trace elements in one product, giving ease of application to the farmers. (b) Several compounds in slow release bioleachable/bioavailable oxide form rather than soluble inorganic salt form. This will enable the trace elements to last for a few seasons rather than leaching just after one season.

(c) Granular form for ease of use and minimises dust during application.

The process of the present invention for the production of EMD, as described hereinabove, is expected to allow for an increase in the efficiency of utilisation of manganese relative to traditional sulphuric acid leach processes for the production of EMD as a result of the novel recycle of process streams.

Modifications and variations such as would be apparent to the skilled addressee are considered to fall within the scope of the present invention.