EMISSION

The present invention concerns the metallic manganese production from manganese ore, in particular from pyrolusite manganese ore. It concerns also CO2 free emission processes.

In the wake of the climate crisis, industries around the world are looking for ways to cut greenhouse gas emissions such as CO2 emissions and the manganese industry is no exception.

However, all known processes of production of metallic manganese from manganese ore generate CO2 since they are carbothermic manganese processes in which carbon dioxide is a by-product of the manganese- producing reactions themselves.

This is in particular the case of the process from the prior art in which an Electrolytic Manganese Metal (EMM) is obtained from pyrolusite ore (Mn as MnCh) through reduction of the ore with coal or oil in a rotary kiln or skinner muffled-hearth furnace followed by sulfuric acid leaching, purification and electrolysis. The reduction is required to turn all the manganese in the ore into a Mn ²⁺ state (MnO) as higher oxidation states (like Mn02, found in the high-grade manganese ore from Gabon and Australia, or the low-grade ore used by the US Bureau of mines) will not be dissolved by acid. The reduction can be accomplished by coal, coke, producer gas, natural gas, butane or oil, each reduction agent containing carbon atom. Therefore such a reduction process generates CO2 according for example to one of the following reactions:

2 MnO ₂ + C => 2 MnO + CO ₂

4 MnO ₂ + CH ₄ => 4 MnO + CO ₂ + 2H ₂ O. The article of Zhang, et al. (Life cycle assessment of electrolytic manganese metal production; Journal of Cleaner Production, 253 (2020), pages 1-11) describes an alternative to the use of pyrolusite ore: the use of carbonated ore (Mn as MnC0 ₃ ). In this case, the ore is partly calcined (as in US 2,392,385) and partly directly sent to leaching. As a consequence, CO2 from the ore is released to the atmosphere at the calcining or at the leaching stage (MnCO ₃ => MnO + CO2). CO2 is also produced by the combustion of the carbonaceous material, usually in the form of coal used in the calcining step.

US 2,766,197 describes a process in which instead of using a calcined ore, manganese slag that is produced by the carbo-thermal reduction of manganese ore in an electric furnace or a blast furnace is used. After crushing, the slag follows the same route as the calcined ore. However the manganese slag used as raw material comes from the reduction of manganese ore by a carbonaceous reductant (coke or coal), which therefore also release CO ₂ in the atmosphere.

In addition to their CO2 emissions, these known processes yield two drawbacks:

- They use coal which brings impurities in the leached solution. Those impurities will need either to be extracted during the hydrometallurgical step, or will end-up in the finished product;

- The EMM electrolysis has a poor faradic yield (60 ~ 65 %) because part of the electricity will be wasted in water's electrolysis that generates hydrogen which is wasted today.

The inventors have surprisingly found that it is possible to substitute the reduction steps with a carbonaceous agent (coal, etc.) by a direct- reduction step using H ₂ gas for non-carbonated manganese ore such as pyrolusite manganese ore. After the reduction, the pre-reduced ore will be leached and follows the same hydrometallurgical steps as the other processes. The present invention therefore concerns a process for recovering metallic manganese from a non-carbonated manganese ore, in particular a pyrolusite manganese ore, without CO ₂ emission, comprising the following successive steps: a) direct reduction step of the non-carbonated manganese ore, in particular the pyrolusite manganese ore MnO ₂ , at a temperature in the range of 600 to 1000°C using H ₂ gas as a reducing agent in order to obtain a pre-reduced ore MnO; b) Leaching of the pre-reduced ore MnO obtained in step a) by an aqueous solution containing sulfuric acid in order to obtain an aqueous solution containing MnSO ₄ ; c) Electrolysis of the aqueous solution containing MnSO ₄ obtained in step b) in order to obtain metallic manganese Mn ⁽⁰⁾ and d) Recovery of metallic manganese Mn ⁽⁰⁾ wherein no carbonaceous reduction agent is used in step a). The main reactions at each step of the process are: a) Direct reduction (in case of the pyrolusite manganese ore): MnO _{2 (s)} + H _{2 (g)} → MnO _(s) + H ₂ O _(g) b) Leaching of the pre-reduced ore by sulfuric acid: MnO _(s) + H ₂ SO _{4 (aq)} → MnSO _{4 (aq)} + H ₂ O c) Electrolysis of the manganese solution: Cathodic reaction: MnSO + 2 e- ð M 2- 4 _(aq) n° _(s) + SO ₄ (aq) Anodic reaction: SO 2- 4 _(aq) + H ₂ O _(l) ð H ₂ SO _{4 (aq)} + 2 e- + ½ O _{2 (g)} In the sense of the present invention non-carbonated manganese ore is intended to mean any manganese ore which does not contain carbonates ore such as MnC0 ₃ .

In an advantageous embodiment, the non-carbonated ore of the invention is a manganese oxide ore such as pyrolusite, romanechite, manganite, and/or hausmannite ore. More advantageously the non-carbonated manganese ore of the invention is the pyrolusite ore MnCh.

In an advantageous embodiment the non-carbonated manganese ore is in the form of lumpy ore, advantageously the ore is run-of-mine, therefore not transformed after its extraction and used as such in the process.

In the sense of the present invention, "lumpy ore" is intended to mean a rocky ore whose particle size distribution is typically between 5 and 75 mm. The particle size is measured by sieving, in particular according to the standard ISO 3310 :1 Part 1 of 2016.

In another embodiment, the non-carbonated manganese ore is crushed, advantageously down to 6.3 mm for the top size, before the reduction step a).

In another embodiment, the non-carbonated manganese ore is sieved, or sieved and crushed, advantageously down to fines with a 6.3 mm top size, before the reduction step a).

If the ore is simply sieved at 6.3 mm, the 6.3 mm sieve passing fraction is used in step a) of the process whereas the 6.3 mm non-passing fraction of ore is used for other applications.

If the ore is submitted to crushing, the ore is first sieved at 6.3 mm and the 6.3 mm non-passing fraction is crushed. The crushed fraction is either blended with the 6.3 mm sieve passing fraction or recycled to the 6.3 mm sieving step. The 6.3 mm sieve passing fraction is used in the reduction step a).

In both case, it results that a nearly or fully 100% passing 6.3 mm fraction of ore is fed to the reduction step a).

The particle size is measured by sieving, in particular according to the standard ISO 3310:1 Part 1 of 2016.

The crushed or non-crushed ore can also be pulverized at a lower dimension in particular at around 100pm for the top size. In this case advantageously step a) is carried out in a fluidized bed on the pulverized ore.

In another embodiment the pulverized ore is pelletized and therefore the non-carbonated manganese ore is in the form of pellets.

If the ore is lumpy, crushed or in form of pellets, the reductions step a) can be carried out in a shaft furnace.

More advantageously, the non-carbonated manganese ore is a lumpy ore.

In another particular embodiment, the process according to the present invention comprises a preliminary step alpha) consisting of drying the ore before the reduction step a). In particular this preliminary step alpha) can be carried out before the crushing and/or sieving and/or pulverization and/or pelletizing of the ore and/or during the crushing and/or sieving and/or pulverization and/or pelletizing of the ore and/or after the crushing and/or sieving and/or pulverization and/or pelletizing of the ore, more advantageously before the crushing and/or sieving and/or pulverization and/or pelletizing of the ore.

The drying step alpha) can be carried out in a separate equipment or naturally or in the upper part of the shaft furnace. Step a) of the process according to the invention consists in the direct reduction step of the non-carbonated manganese ore, in particular the pyrolusite manganese ore MnO ₂ , using H ₂ gas as a reducing agent in order to obtain a pre-reduced ore MnO.

Indeed when H ₂ is used as the reducing agent, it has been surprisingly found by the inventors that the reduction step a) is faster and the reaction is more complete than when CO is used as the reducing agent. Therefore there is a reduction in the residence time of step a) and an improvement in the leaching yield of step b).

Step a) is therefore carried out at a temperature in the range of 600- 1000°C, advantageously at a temperature above 750°C to avoid reoxidation of the ore, even still more advantageously at a temperature >850°C (for example in the range of 850-1000°C, more advantageously in the range of 850 - 900°C). Indeed it has been surprisingly found that at a temperature >850°C, in particular 850-900°C, there is no re-oxidation of the ore after the reduction step and the reaction kinetics is particularly interesting. At a temperature above 1000°C there is a risk of sintering the ore and the energy balance is not interesting. In particular step a) lasts 2 to 5 hours, advantageously 4 hours.

In the process according to the invention no carbonaceous reduction agent (such as coke or coal) is used in step a). H ₂ can be used as the only reducing agent or can be used in combination with other non- carbonaceous reduction agent such as ammoniac (NH ₄ ), more advantageously H ₂ is the only reducing agent.

Step a) can be carried out at atmospheric pressure. It is in particular not carried out at reduced pressure.

Step a) can be carried out under oxygen-free atmosphere such as a mixture of H ₂ and N ₂ . In particular, no air is present to avoid any problem of explosiveness of the gas mixture. Step a) can be carried out in a shaft furnace or a fluidized bed, more advantageously in a shaft furnace.

Advantageously the H ₂ gas used in step a) is water free. Indeed water vapor can have a strong inhibiting effect on the reduction reaction, contrary to what could be expected from the thermodynamics. The retardation is probably due to adsorption of water on active reaction sites at the solid-gas reaction interface.

Step a) can be operated in continuous mode or batch-wise. Advantageously it is operated in continuous mode.

Step b) of the process according to the invention consists in the leaching of the pre-reduced ore MnO obtained in step a) by an aqueous solution containing sulfuric acid (H ₂ SO ₄ ) in order to obtain an aqueous solution containing MnSO ₄ .

More advantageously the anolyte of step c) is cycled back for the leaching step.

In particular the initial pH of the leaching step is in the range of 1-2.

Advantageously the final pH of the leaching step is in the range of 5.2-7. More advantageously this final pH is obtained by neutralization with MnO to precipitate impurities.

Advantageously step b) is carried out at ambient temperature.

If step a) is carried out in a shaft furnace and the ore used in step a) is not in the form of pellets, the pre-reduced ore MnO obtained in step a) can be crushed or pulverized before the leaching step b). The leaching step b) is carried out by methods well known by the one skilled in the art such as described in US 2,766,197.

Step c) of the process according to the invention consists in the electrolysis of the aqueous solution containing MnSO ₄ obtained in step b) (leachate solution) in order to obtain metallic manganese Mn ⁽⁰⁾ . The electrolysis step c) is carried out by methods well known by the one skilled in the art such as described in US 2,766,197.

In particular the anode is formed of lead alloyed with silver. The cathode can be formed of stainless steel and the voltage between the anode and cathode is ~ 5 V. A porous diaphragm separates the anode compartment from the cathode compartment.

Advantageously the electrolyte circulating through the cathode compartment consists of the aqueous solution containing MnSO ₄ obtained in step b). The electrolyte circulating through the cathode consists typically of 12-15 g/L of Mn ²⁺ , 135-150 g/L (NH ₄ ) ₂ SO ₄ . In particular the electrolyte circulating through the anode compartment consists of 12- 15 g/L Mn ²⁺ in 125 g/L (NH ₄ ) ₂ SO ₄ and 30- 35 g/L H ₂ SO ₄ .

The pH in the cathode compartment is advantageously maintained around ~7. The pH in the anodic compartment in the range of 7 - 8. The temperature is typically 35-40 °C.

Advantageously the process according to the invention contains an intermediate step bl) of purification and filtration of the leachate obtained in step b) (therefore of the aqueous solution containing MnSO ₄ ), and step c) is carried out on the filtrate obtained in step bl). Indeed sometimes impurities such as nickel, cobalt, zinc, lead, iron, arsenic and/or antimony are present in the leachate and need to be removed, for example by precipitation in particular using hydrogen sulfide H ₂ S or sodium sulfide Na ₂ S or ammonium sulfide (NH ₄ ) ₂ S, after removal of iron by standard methods, such as oxidation to ferric state and precipitation at a suitable pH, as disclosed in US2259418 and US2325723. Finally step d) consists in the recovery of metallic manganese Mn ⁽⁰⁾ , in particular by stripping the metal solid deposited on the cathode. In an advantageous embodiment the process according to the present invention comprises a preliminary step alpha) before step a) which consists in the hydrogen production by water electrolysis. Indeed in this case even the production of H ₂ is CO ₂ -free. The H ₂ production of step alpha) can be carried out in in a separate unit, or in situ. Advantageously it is an in-situ step. Step alpha) can use alkaline electrolysis in particular using Ni catalyst, Proton Exchange Membrane (PEM) (in particular using platinum iridium catalyzer) or Solid Oxide Electrolysis Cells (SOEC), advantageously an alkaline or proton exchange membrane technology. These technologies are well known by the one skilled in the art. In an advantageous embodiment, the process according to the present invention comprises an intermediate step c1) after step c) of recovery of H ₂ by-product obtained during the electrolysis step c). Indeed in addition to the above-mentioned main reactions, the electrolysis generates side reactions leading to unwanted hydrogen and manganese dioxide production: Cathodic side reaction: H S - 2- 2 O _{4 (aq)} + 2 e ð H _{2 (g)} + SO ₄ (aq) Anodic side reaction: MnSO + SO 2- + 2.H O ð MnO - 4 _{(aq) 4 (aq) 2 (l) 2} + 2.H ₂ SO _{4 (aq)} + 2 e _. Step c1) can be carried out by enclosing the cathode of step c) in a sealed container in order to recover only H ₂ or by using a gas suction mean coupled to gas separation unit in order to separate H ₂ from the other gas. In an advantageous embodiment the H ₂ by-product obtained in step cl) is used in step a) of the process according to the present invention.

Indeed the most common way to improve the faradic yield of step c) and decrease the side reactions is to add selenium compounds such as selenium oxide (SeO ₂ ). The gain is around 10 %, but causes environmental issues.

However, if the hydrogen produced during the Mn electrolysis of step c) is recovered in step cl) and used for the direct reduction of the manganese ore of step a), this will improve the global energy efficiency of the process according to the present invention without the need of using selenium compounds. Indeed the low faradic yield of the Mn electrolysis will not be wasted but valorized internally as hydrogen. Therefore advantageously step c) is a selenium-free electrolysis step. The process according to the invention will indeed make the Se-free electrolysis more attractive as the lower faradic yield will have less impact on the profitability.

Moreover advantageously the dedicated water hydrolysis unit of step alpha) may even be removed, or at least strongly downsized if the process according to the present invention contains step cl).

In an advantageous embodiment the electricity used in the process according to the invention is a CO ₂ -free electricity such as nuclear electricity or electricity from renewable energy such as solar electricity, wind electricity, hydroelectricity and marine electricity. In this particular embodiment the process according to the present invention does not at all produce CO ₂ .

The present invention will be better understood in reference to the description of the examples and figures below. In the examples, it is understood that MnOx is a formula standing for all kind of manganese oxides naturally present in the Mn ores. “x” is therefore an indicator of the oxidation state of Mn, when x=1 it means that any Mn present in the ore has an oxidation state of +II whereas when x=2 it means that any Mn present in the ore has an oxidation state of +IV ( _e.g. x = 1.0 for MnO, 1.5 for Mn ₂ O ₃ , 2.0 for MnO ₂ ). Figure 1 represents the heating rise followed by the reduction step of Ex 1 (temperature as a function of time) and the type of gas introduced in the furnace at each step according to example 1. Figure 2 represents “x” as an indicator of the global oxidation degree of the manganese (MnO _x ) as a function of time in hours in the tubular furnace of Example 1 for Ex1, Comp Ex1 and Comp Ex2. Figure 3 represents “x” as an indicator of the global oxidation degree of the manganese (MnO _x ) as a function of the temperature in °C in the tubular furnace of Example 1 for Ex1, Comp Ex1 and Comp Ex2. Figure 4 represents the reduction rate (dx/dt) as a function of time in hours in the tubular furnace of Example 1 for Ex1, Comp Ex1 and Comp Ex2. Figure 5 represents “x” as an indicator of the global oxidation degree of the manganese (MnO _x ) as a function of time in hours in the tubular furnace of Example 2 for Ex2 and Ex3. Figure 6 represents “x” as an indicator of the global oxidation degree of the manganese (MnO _x ) as a function of time in hours in the tubular furnace of Example 1 and Example 2 for Ex1 and Ex3. Figure 7 represents the leaching yield in % of the leaching step carried out according to example 3 for Ex1, Comp Ex 1 and Comp Ex2. Example 1: direct reduction of ore using H ₂ as the reducing agent (Ex 1) and comparison with reduction using CO as the reducing agent (Comp Ex 1) or a mixture of CO and H ₂ as the reducing agent (Comp Ex 2).

In a horizontal tubular furnace, a pyrolusite manganese ore (1.3 kg) crushed to below 6.3 mm with a jaw crusher is placed at 25°C in the reaction chamber and heated under controlled atmosphere (N ₂ , CO and/or H ₂ ). For Comp Ex 2 the mixture consists of 50/50 by volume of CO/H ₂ . The final temperature is 850°C, the gas flow is 300 l/h, the heating rate is 5°C/min with a temperature stage at 200°C. The heating rise is presented in figure 1.

The results are illustrated in figures 2, 3 and 4.

As can be seen in these figures, the reduction reaction using CO as the reducing agent (comp Ex 1) starts earlier than the other ones using H ₂ (Ex 1) and a mixture of H ₂ and CO (Comp Ex 2). However the reduction reaction using H ₂ as the reducing agent (Ex 1) is faster and the full reduction to MnO is reached earlier. The reduction reaction using a mixture of CO/H ₂ as the reducing agent (Comp Ex 2) takes from both: earlier start than Exl & faster kinetics than Comp Exl. There is no obvious difference in SEM (Scanning Electron Microscopy) observations of the MnO obtained between the use of CO or H ₂ as the reductant.

Example 2: direct reduction of ore using H ₂ with lumpy (Ex 2) or crushed ore (Ex 3)

The process is the same as described in Example 1 except that the final temperature is 700°C and the ore is either lumpy (ex 2) or crushed below 6.3mm (Ex 3) and the heating rate is 2°C/min. The results are illustrated in figure 5: it can be seen that the finer the ore size, the faster the reduction reaction. Moreover when comparing the temperature rate between Ex3 and Ex 1 of example 1, it can be seen that faster heating leads to faster reduction (Figure 6).

Example 3: Leaching of MnO in Electrolytic Manganese Metal production conditions followed by electrolysis to obtain Mn° Exl, Comp Exl and Comp Ex2 obtained according to example 1 are leached by synthetic anolyte to mimic EMM production conditions: The leachate solution obtained has the following composition:

125 g/L (NH ₄ ) ₂ SO ₄ + 40 g/L H ₂ SO ₄ + 12 g/L Mn

The leaching step at an initial pH of 1.5 is followed by neutralization with MnO at a pH of 6.5-7, final purification with Na ₂ S and electrolysis using a 10cm ² cathode.

The leaching kinetics is fast (~2 h) for all samples. There are no major differences.

The leaching yield is illustrated in figure 7 and shows a better yield for the examples according to the invention than for Comp EXI (CO as the reducing agent).

23g of Mn are obtained on the cathode (70% current efficiency).