Manganese Metallurgy Part I

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(2007) 137–159www.elsevier.com/locate/hydromet

Hydrometallurgy 89

Manganese metallurgy review. Part I: Leaching of ores/secondarymaterials and recovery of electrolytic/chemical manganese dioxide

Wensheng Zhang ⁎, Chu Yong Cheng

Parker Centre for Integrated Hydrometallurgy Solutions, CSIRO Minerals, PO Box 7229, Karawara, WA 6152, Australia

Received 6 July 2007; received in revised form 24 August 2007; accepted 25 August 2007Available online 1 September 2007

Abstract

The world rapidly growing demand for manganese has made it increasingly important to develop processes for economical recoveryof manganese from low grade manganese ores and other secondary sources. Part I of this review outlines metallurgical processes formanganese production from various resources, particularly focusing on recent developments in direct hydrometallurgical leaching andrecovery processes to identify potential sources of manganese and products which can be economically produced.

High grade manganese ores (N40%) are typically processed into suitable metallic alloy forms by pyrometallurgical processes.Low grade manganese ores (b40%) are conventionally processed by pyrometallurgical reductive roasting or melting followed byhydrometallurgical processing for production of chemical manganese dioxide (CMD), electrolytic manganese (EM) or electrolyticmanganese dioxide (EMD).

Various direct reductive leaching processes have been studied and developed for processing low manganese ores and oceanmanganese nodules, including leaching with ferrous iron, sulfur dioxide, cuprous copper, hydrogen peroxide, nitrous acid, organicreductants, and bio- and electro-reductions. Among these processes, the leaching with cheap sulfur dioxide or ferrous ion is mostpromising and has been operated in a pilot scale. The crucial issue is the purification of leach liquors and the selective recovery ofcopper, nickel and cobalt is often difficult from solutions containing soluble iron and manganese. For treatment of manganesebearing materials including waste batteries, spent electrodes, sludges, slags and spent catalysts, a leaching or reductive leachingstep is generally needed followed by various purification steps, which makes the processes less economically viable.

It is concluded that the recovery of manganese from nickel laterite process effluents which contain 1–5 g/L Mn offers a growinglow cost resource of manganese. Part II of this review considers the application of various solvent extraction reagents andprecipitation methods for treating such manganese liquors.© 2007 Elsevier B.V. All rights reserved.

Keywords: Manganese; Metallurgy; Leaching; Chemical manganese dioxide; Electrolytic manganese dioxide

1. Introduction

Manganese is an important metal in human life andindustry. In recent years, the world manganese demandhas been driven by soaring steel production, particularly

⁎ Corresponding author.E-mail address: [email protected] (W. Zhang).

0304-386X/$ - see front matter © 2007 Elsevier B.V. All rights reserved.doi:10.1016/j.hydromet.2007.08.010

in China. Steelmaking, including its iron making com-ponent, has accounted for most of the world manganeseconsumption, presently in the range of 85% to 90% ofthe total demand (Corathers, 2004, 2005). Most of themanganese is consumed in steelmaking in the form ofmanganese ferro-alloys. In 2004, manganese consump-tion was 60% higher than that of 2003. World pro-duction of manganese ore rose steadily by 9% in 2004

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138 W. Zhang, C.Y. Cheng / Hydrometallurgy 89 (2007) 137–159

and 2005 on a contained-weight basis compared withthat in 2004. The price of metallurgical-grade ore(N40%) increased by 16% in 2004 and 63% in 2005internationally. Combined world production of ferro-manganese and silicomanganese rose by 8% on a gross-weight basis compared with that in previous year(Corathers, 2004, 2005).

The rapid growing demand for manganese metal asan alloying element comes primarily from the alumin-ium industry followed by the steel industry and that forelectrolytic manganese dioxide (EMD) and chemicalmanganese dioxide (CMD) come from the primary andsecondary battery industries. Industry EMD usage inalkaline batteries in 2002 exceeded 230,000 t per annumwith an annual growth rate in excess of 9.6% between1996 and 2002 (Anon, 2003). Future growth in theEMD market is expected to exceed current growth rates,as global demand for batteries will increase in areassuch as mobile communications. A wide range of newproducts, such as disposable mobile telephone batteries,could see growth rates of up to 8% a year (Anon, 2003).Electrically powered consumer products, from camerasand watches to sophisticated children's games and toysare also an area of robust growth. Extra growth in thedemand for EMD is also expected from the secondary orrechargeable battery market as lithium manganese ionbatteries (that use EMD in their manufacture) replacecobalt type batteries as the cheaper and more environ-mentally friendly alternative. Growth is also likely tocome from the use of EMD in hybrid electric vehicles(HEV's). As the environmental impact of traditionalcars continues to cause problems in densely populatedcities, the increased usage of HEVs to minimize pol-lution levels seems inevitable.

With the rapid growing demand for manganese prod-ucts, low grade manganese ores (b40%), manganeseocean nodules and secondary manganese sources, whichcannot be economically processed by conventional py-rometallurgical processes or pyro-pre-treatment, be-come increasingly important sources of manganese. Inrecent years, various hydrometallurgical processes havebeen studied and developed for recovery of manganesefrom these manganese sources.

In this part of review, conventional pyrometal-lurgical processes and pyro-pre-treatments for produc-tion of major manganese products are briefly outlined.Recent developments in direct hydrometallurgicalprocesses for leaching and recovery of manganese arereviewed in details.

The aims of this part review are to obtain a generalknowledge of manganese metallurgy including pyro-metallurgy, pyro-hydrometallurgy and hydrometallurgy

and to identify potential sources of manganese andproducts which can be economically produced.

2. Manganese resources and pyro-processes

2.1. Manganese resources

Manganese is the twelfth most abundant element inthe earth's crust (0.096%). Its deposits are generally ofsedimentary origin, with oxide ore layers inter-beddedwith iron-rich formations (Kaneko et al., 1993). Themost common mineral is pyrolusite, which is mainlyMnO2. Manganese is also found in several minerals,such as pink rhodochrosite (MnCO3), rhodonite(MnSiO3), black manganite (MnO(OH)), and alabandite(MnS) (Calvert, 2004). The main sources of manganesecome from the former U.S.S.R, Brazil, South Africa,Australia, Gabon and India. Russia and South Africaproduce about 85% of the world's pyrolusite.

Manganese nodules or ferro-manganese concretions,usually containing 30–36% Mn, have been found onocean floors (Calvert, 2004) and could provide anothersource of manganese. These nodules are found in boththe Atlantic and Pacific Oceans, but principally in thePacific Ocean. Although the primary interests in deep-sea nodules are nickel, copper, and cobalt values, thelarge quantities of manganese could also be of futureimportance. As a result, much research work has beendevoted to recovering not only nickel, copper and cobaltbut also manganese as well.

Manganese bearing materials such as waste batteriesand spent electrodes, spent catalysts, steel scraps, sludgeand slag are secondary manganese sources. Manganeseminerals are often associated with zinc sphalerite oresand nickel laterite ores, which are leached and rejectedto the waste effluents in subsequent processing steps.These manganese-containing industrial waste effluentscould be potentially important manganese sources. Forinstance, in the leach solution of the Murrin MurrinOperation of Minara Resources Ltd., the ratio of nickel/manganese concentrations is in the order of 2:1. Itscurrent nickel production is 36,000 t per annum.Therefore, the manganese in the waste stream is about18,000 t per annum (Zhang and Cheng, 2006).

2.2. Manganese pyro-metallurgy

Metallurgical grade manganese ores (N40% Mn) areusually processed into suitable metallic alloy forms bypyrometallurgical processes, which are very similar to ironpyrometallurgical processes. Ferro-manganese is a com-monly used alloy. In its production process, a mixture of

139W. Zhang, C.Y. Cheng / Hydrometallurgy 89 (2007) 137–159

manganese ore, reductant (a form of carbon) and flux(CaO) are smelted at above 1200 °C to enable reductionreactions and alloy formation. Standard grade ferroman-ganese contains more than 76%Mn and about 7% C, andcan be produced either in a blast furnace or in an electricfurnace (Yucel and Ari, 2001). High carbon ferro-manganese can be made by three different practices:blast furnace, discard slag electric furnace, and highmanganese slag electric furnace. Medium carbon ferro-manganese can be produced by a de-carbonation processor through a redox reaction between silicon in thesilicomanganese alloy and manganese ores. Low carbonferro-manganese is produced by the reaction of manga-nese ore and low carbon silico-manganese. Silico-manganese is another commonly used alloy, containing14–16% Si, 65–68% Mn, and about 2% carbon. Lowercarbon levels result when the silicon content is increased.Special grades with up to 30% Si are produced for use inthe manufacture of stainless steel. The production ofsilico-manganese is very similar to that of electricsmelting of ferro-manganese. It differs in furnace chargewhich contains large amount of quartz (SiO2) (Yucel andAri, 2001). Off-grade silicon bearing materials such asferrosilicon fines, drosses, or ladle dig-outs provide arelatively cheap source of silicon for the silico-manganesemaking operation because re-melting these materials usesless energy than smelting silicon.

2.3. Manganese pyro-hydrometallurgy

A pyrometallurgical pre-treatment followed by hydro-metallurgical processing plays an important role in thetreatment of lowgrademanganese ores andmanganese seanodules containing Ni, Co and Cu values as their oxides.These metal oxides in the nodules often occur in thelattices of iron and manganese minerals. Therefore,breaking up these lattices by pyrometallurgical reductionor hydrometallurgical reductive dissolution is an essentialstep for the satisfactory recovery of valuable metals. Thesuggested pyrometallurgical pre-treatments are smelting(Cooper et al., 1959), reduction-roasting (Rolf, 1969),sulphatising (Freitas et al., 1993), and chloridising (Cooperet al., 1959). Compared with pure hydrometallurgicalprocesses, combining pyro-hydrometallurgical treatmentsyield better results for efficient recovery of Ni, Co, Cu andMn values from poly-metallic manganese nodules (Kohgaet al., 1995), but they require more energy consumption.

In sulphation roasting, the manganese ores or Mn-bearing materials are roasted in the presence of sulfuricacid or ammonium sulfate to convert the manganeseminerals to soluble sulfates. Sulphation of manganese orewith gaseous SO2 has been reported (Freitas et al., 1993).

In this case, the gaseous SO2 plays dual roles in thesulphation process: a reductant and a sulphation agent.Sulphation roasting followed by water leaching has beeninvestigated for recycling of zinc-carbon spent batterieswith a production of manganese and zinc sulfates (Abbaset al., 1999). The process involves mechanical separationand sulfuric acid leaching, sulphation roasting in thepresence of sulfuric acid or ammonium sulfate, followedby water leaching. Maximum manganese recoveryefficiency (93.4%) is achieved by using an ammoniumsulfate roasting process. In a patented process byKasimovet al. (1990), the Mn-containing slag, after mixingwith (NH4)2SO4, was roasted and then leached withaqueous (NH4)2SO4 to recover manganese compound atlow (NH4)2SO4 consumption. The sulfate-containingresidue was then pressure-leached at 100–150 °C withaqueous (NH4)2CO3 at a NH3:CO2 mole ratio of 1:1–1.5 and heated to recover (NH4)2SO4 for recycling. Thisapproach for recycling (NH4)2SO4 appears to be ex-pensive involving NH3 input and high pressure leachstep, and thus less practical.

Smelting or reduction roasting at about 700–900 °Cfollowed by sulfuric acid leaching is by far the mostcommonly employed method in the manganese industryfor production of intermediate or final manganoussulfate for further electrowinning process (Paixao etal., 1995). The reduction reaction can be presented as:

MnO2 þ CO=H2→MnO þ CO2=H2O ð1Þ

MnO2 þ C→MnO þ COðCO2Þ ð2Þ

This process converts higher valent manganese oxidesto lower ones which are readily soluble in sulfuric acid.

3. Leaching processes

The typical leaching processes involve a chemicalreductive step, including bio-leaching and electro-leaching except for acid leaching of Mn(II) ores. Puri-fication, separation and final recovery processes such assolvent extraction, electrolysis and electrowinning, andother recovery processes are then used.

3.1. Reductive leaching with ferrous iron solution

Treatment of manganese ore or manganese bearingslimes with acidified ferrous sulfate or pickle liquors hasbeen reported in several studies (Brantley and Rampa-cek, 1968; Das et al., 1982; Vu et al., 2005).

A rather novel process for recovery of manganesewith pickle liquor (weak FeSO4-H2SO4-(NH4)2SO4

6

Fig. 1. Flowsheet for the process of Mn recovery using pickle liquor to leach ore (based on Brantley and Rampacek 1968).


solutions) from low grade manganese ores was devel-oped by US Bureau of Mine (Brantley and Rampacek(1968)). The process is schematically shown in Fig. 1.This process features the simultaneous high temperaturehydrolysis of Fe(III) and precipitation of manganeseas manganese double salt (NH4)2Mn2(SO4)3 which isthen separated from coprecipitated Fe2O3 by disso-lution of the salt in water. A relative pure manganesepregnant solution for electrolysis is expected to beobtained in this way. The economical concern of thisprocess is its involvement of expensive H2O2 oxidantand a high temperature autoclave process. However,recent developments in pressure oxidation and auto-clave technologies may make this approach moreattractive.

A study by Das et al. (1982) showed that the reactionof MnO2 in the low manganese ore with ferrous sulfatecould occur in three ways:

with neutral ferrous sulfate solution

MnO2 þ 2FeSO4 þ 2H2O¼ MnSO4 þ FeðOHÞSO4 þ FeðOHÞ3 ð3Þ

with ferrous sulfate solution and small amount of acid

MnO2 þ 2FeSO4 þ H2SO4

¼ MnSO4 þ 2FeðOHÞSO4 ð4Þand with ferrous sulfate solution and excess amount ofacid

MnO2 þ 2FeSO4 þ 2H2SO4

¼ MnSO4 þ FeðSO4Þ3 þ 2H2O ð5Þ

More than 90% of manganese could be extracted byleaching the low manganese ore with a stoichiometricamount of ferrous sulfate at 90 °C for 1 h with solid-liquid ratio of 1:10. The leach slurry obtained under theabove conditions was gelatinous and difficult to filter.An addition of sulfuric acid made filtration easier andgreater Mn extraction, but resulted in 3- 4-fold more ironextraction (Das et al., 1982).

Dundua and Agniashvili (1999) used FeSO4 asleaching agent for manganese recovery from residualslimes in the electrochemical production of manganesedioxide. The resulting liquor was purified by removal ofFe(II) ions using manganese carbonate from low-gradeores. The optimised leachingwas proposed formanganeserecovery from sludge wastes in the closed-cycle electro-chemical production of MnO2.

3.2. Reductive leaching with sulfur dioxide or sulfitesolutions

Reductive leaching of manganese nodules and lowgrade manganese oxide ores with SO2 dioxide or sulfitesalts has well been documented in the literature. It is aneffective reductant for higher manganese oxide mineralssuch as MnO2 and manganese nodules in the hydro-metallurgical leaching process. Aqueous SO2 has beeninvestigated for both percolation and agitation leaching.

Typical studies on leaching of manganese ores withSO2 or sulfite salts are summarised in Table 1.

The reactions which may occur during leachingmanganese oxides are (Petrie, 1995):

MnO2 þ SO2→Mn2þ þ SO2−4 ð6Þ

MnO2 þ 2SO2→Mn2þ þ S2O2− ð7Þ


2MnOOH þ SO2 þ 2Hþ→2Mn2þ þ SO2−4 þ 2H2O

ð8Þ2MnOOH þ SO2 þ 2Hþ→2Mn2þ þ S2O

2−6 þ 2H2O

ð9ÞThe SO2 is oxidised to SO4

2− and a proportion ofdithionate S2O6

2− by-product depending on solution pH,temperature and redox potentials (Das et al., 1998;Ward et al., 2004; Ward, 2005). A radical mechanismwas proposed for formation of dithionate, involving theformation of SO3 radicals and dimerisation to S2O6

2−

(Petrie, 1995).In early 1946, a dithionate process was first developed

for recovery of manganese from low-grade ores (Ravitzet al., 1946). This is essentially a SO2 reductive process.The advantage of this process is to utilise the dithionateintermediate product of the reductive reaction to stabiliseboth the reduced Mn(II) and calcium in solution whileextra calcium is precipitated as CaSO4 precipitate. Thisallows lime to be used as a cheap neutralisation agent. Inthe semi pilot-plant tests, about 40 runs were carried out onores assaying 13–18%Mn.Manganese recoverywasmorethan 90% in most cases. A final product containing morethan 60% Mn, meeting ferro-grade specifications, wasproduced. For each unit of manganese extracted, 1.5–2.2unit of SO2 and 1.1–1.5 CaO were consumed. Thequantity of SO2 consumedwas 60–70%of that introduced.

Dithionate was also used for the treatment of low-grademanganese slimes (Maslenitskii et al., 1969) at alaboratory scale. A recovery of 90% Mn was achievedwhen the carbonate manganese accounted for 50% of thetotal manganese content. The manganese content in the

Table 1Reductive leaching of Mn ores/slimes by aqueous SO2/sulfite

Feed Reagents Conditions

MnO2 minerals Aq. SO2 Studied leaching chemistry

Rich Mn ore (N40% Mn) Aq. SO2 Atm. pressure and room temLow Mn ore (b40% Mn) Dithionate Lime for neutralisation

Aq. SO2 Atm. pressure and room tem

Aq. SO2 b60 °C controlledEh for adding SO2

Low Mn slimes Dithionate –

Fe–Mn concretions (NH4)2SO3–bisulfite–H2SO4

pH 3.5–5.5 30–70 °C

Aq. SO2 Kinetic study

Low Mn ores (Percolation) Aq. SO2 Size 1.3–2.5 cm 1.5%-sat.[SO2] 15–20 days

produced concentratewas always 55–60%, irrespective ofthe content in the slimes. The concentrates could be usedto produce ferromanganese by the fluxless method or tomake electrolytic manganese with the dithionate process.

However, it is generally undesirable to form dithionatein SO2 leach processes as it requires long residence time tooxidise the dithionate to sulfate and would contaminatethe manganese sulfate products by slowly reacting torelease sulfur dioxide. A process that minimises theformation of (S2O6

2−) to below 1 g/L by addition of SO2

continuously at a controlled rate determined by monitor-ing the solution potential has been developed and reported(Ward et al., 2004; Ward, 2005, see also Section 5.4).

Direct aqueous SO2 leaching was used for rich Greekmanganese ores (Grimanelis et al., 1992). The reactionwas found fast with N95% Mn recovery in about 10 minand with low iron impurity. In the studies using lowmanganese ores (Naik et al., 2000, 2002, 2003), the ratewas proposed to be controlled by SO2 diffusion to thereaction surface, evidenced by a low apparent activationenergy of 16.6 kJ/mol. Twice the stoichiometric quantityof SO2 was required for dissolution of manganese.The (NH4)2SO3–bisulfite mixture at the bisulfite/H2SO4

(w/w) ratio of 1:(0.52–1.55) was also used as the reduc-tant in leaching the Fe–Mn ore concretions (Sventsitskiiet al., 2003; Partenov et al., 2004). The reaction ratewas found to be first order with respect to SO2 concen-tration and controlled by diffusion of SO2 to the reactionsurface (Partenov et al., 2004).

In processing manganiferrous silver ores where man-ganese and silver minerals are intimately associated, thereductive leaching by aqueous SO2 extracts manganeseinto solution, leaving silver minerals in the residue

Key results Ref

Sulfate/dithionate ratiodepends on pH, Eh and Tem.

Petrie 1995 Das et al. 1998

p. N95% Mn in 10 min Grimanelis et al. 1992N60% Mn, meetingferro-manganese

Ravitz et al. 1946

p. Rapid leaching rate95% Mn recovery

Naik et al. 2000; 2003

Minimised dithionate to 1 g/L Ward et al., 2004; Ward 2005

90% Mn recovery55–60% Mn content

Maslenitskii et al. 1969

96% Mn recovery Sventsitskii et al. 2003

E =16,6 kJ/mol SO2 diffusioncontrol First order: [SO2]

Partenov et al. 2004

Rate depends on [SO2]and pH N90% Mn recovery

Abbruzzese 1987 Pahlman andKhalafalla 1988 Abbruzzese 1990


which could be recovered by subsequent cyanidation(Pahlman et al., 1987).

In percolation (Abbruzzese, 1990), the extent ofthe extraction depends mainly on the dissolved SO2

concentration and the acidity of the leach solution, whichgovern the thermodynamic conditions of the system.The pH must be maintained below the pKa and theconcentration of SO2 around the saturation value (6%).Ambient temperature and small particle size favour man-ganese recovery. The maximum extraction was achievedin 15–20 days of leaching. Simultaneous dissolution ofiron was found to be significant when the ore containedgoethite and siderite. Dissolution of calcium could beinhibited by low flow rate of the leach solution. Over 90%Mn recovery could be achieved (Pahlman and Khalafalla,1988). Pilot percolation plant leaching tests (Abbruzzese1987) showed that total manganese recovery was 94.4%whichwas slightly lower than values obtained in the batchtests. The pregnant solution of MnSO4 containing mainlyFe and Zn impurities was purified using 20% DEHPAin kerosene for extraction of Fe(III) at pH 1.5–1.8 andZn at pH 4. The main concerns on the application ofthe aqueous SO2 for leaching of manganese ores inpercolation is the control of SO2 and the acidity, whichappear difficult and costly.

3.3. Reductive leaching with organic reductants

Many studies have focused on reductive leaching ofmanganiferrous ores containing tetravalent manganeseusing organic reductants, including sawdust (Sanigokand Bayramoglu 1988), glucose, sucrose (Veglio andToro 1994a,b), lactose (Ali et al., 2002 Ismail et al.,2004), glycerine (Arsent'ev et al., 1991), oxalic acid,

Table 2Reductive leaching by organic reductants

Reductant Ore Conditions

Glucose Mn nodules 2.5 M NH3, 0.37 M NH4Cl, 0.2glucose/g of nodules, 85 °C, 4

Sawdust (C6H10O5)n Low Mn Ore Aq. H2SO4

Sucrose (C12H24O11) Pyrolusite 50–90 °C 20 g/L of sucrose H2

Mineral=0.98 Sucrose:MineralLactose (C12H22O11) Low Mn Ore 100 mesh, 2 h, 90 °C H2SO4/

MnO2 =1.8 lactose/Mn-ore=0.Mn–Ag Ores Two stage leach with H2SO4

and then thioureaGlycerine Mn ores Aq. H2SO4 glycerine/MnO2=(1Triethanolamineand Thiosulfate

MnO2 H2SO4:Mn b2.6 Thiosulfate/MnO2=(0.25–0.28)

Oxalic Acid (OX) Low Mn Ore 30.6 g/l OX, 0.5 M H2SO4 85 °Carboxylic Acid(15%) HF (9%)

Ferro columbite HF+tartaric (TR), citric,formic, oxalic (OX) acids

citric acid, tartaric acid, formic acid (Sahoo et al., 2001;Rodriguez et al., 2004), and triethanolamine andthiosulfate (Yavorskaya et al., 1992). The conditionsand main results are summarised in Table 2.

The carbohydrate reductants are considered to becost effective and non-hazardous, which may be usedeither in pure form or from industrial wastes (Veglioand Toro 1994a,b). The stoichiometry was proposed asfollows.

C6H12O6 þ 12MnO2 þ 24Hþ

¼ 6CO2 þ 12Mn2þ þ 18H2O ð10ÞThe reaction products of the glucose were identified

using HPLC (Furlani et al., 2006). In sulfuric acid me-dium, the glucose was oxidized via formation of mono-carboxylic polyhydroxyacids and formic acid. Formicacid was identified and quantified as the major com-ponent. Poly-hydroxyacids, especially glyceric and gly-colic acid, were found in minor quantities together withtraces of gluconic acid.

Sawdust and oxalic acid are also of interest toindustrial applications due to its cheap cost andavailability.

3.4. Bio-reductive leaching

In addition to the studies on chemical leachingprocesses, a certain emphasis has been placed onreductive biological treatment that generally uses hetero-trophic microorganisms (Abbruzzese et al., 1990; Veglioet al., 1997). Themechanism can be direct and indirect. Indirect leaching mechanism, the bacteria are capable ofutilizing the MnO2 as a final acceptor of electrons in the

Key Results References

gh

(%) Cu 100, Ni 90, Co 60 Das et al. 1986

90–95% Mn (99.6% pure) Sanigok andBayramoglu 1988

SO4:=0.1

94–95% Mn recovery Veglio and Toro1994a; 1994b

7590–92% Mn recovery Ali et al. 2002

Ismail et al. 200497% Mn 98% Ag and Au Ziyadanogullari and

Buyuksahin 1995–2.5) Increased Mn recovery Arsent'ev et al. 1991

Maximum Mn recovery Yavorskaya et al. 1992

C, 105 min. 98.4% Mn recovery Sahoo et al. 2001TR increases %Fe andMn OX decreases %Mn

Rodriguez et al. 2004


respiratory chain of their metabolism, instead of oxygen.In the indirect mechanism, the reductive process is as-sociated with the formation of reductive compounds,resulting from their metabolism. In all cases, the bio-logical process occurs in the presence of organic carbonand energy sources.

The mechanism of the bioleaching was found to bepredominantly indirect through the production of organicacids, mainly oxalic acid and citric acid in the leachingmedium which reduced manganese oxides (Acharya etal., 2003).Oxalic acid is potentially of greater interest thancitric acid due to its reductive activity forMnO2 accordingto the following reaction (Abbruzzese et al., 1990):

MnO2 þ 2Hþ þ C2O4H2 ¼ Mn2þ þ 2 CO2 þ 2H2O

ð11Þ

The manganese ions (released as manganese oxalateand manganese citrate) formed a product layer onthe ore, and the reaction was proposed to be controlledby diffusion of the reactants through the permeableproduct layer (Acharya et al., 2003). The kineticequation could be expressed by the shrinking coremodel [1− (1−α)1/3]2 kt (where α is the fraction ofmanganese reacted, k the rate constant and t the reactiontime) (Acharya et al., 2003).

Acharya et al. (2001) investigated microbial extrac-tion of manganese from the low grade manganese orewhich was collected from Joda East manganese minesof Orissa. The microorganism used was an indigenousfungal culture, Penicillium citrinum, isolated from thetop soil of the same mines. An extraction of 64% man-ganese was obtained within 30 days.

Research workers in Greece and Italy (Veglio et al.1997) investigated a batch and semi-continuous processfor bioleaching of manganiferrous minerals with hetero-trophic mixed microorganisms. 95–100% of manganeseextraction could be obtained in 36–48 h of treatmentusing a content of pulp containing 20% (w/v) of orehaving a Mn grade of 17–20%. Final tests in a pilotplant (V=70 L) were performed under non-sterilisedconditions. The experimental results showed a technicalfeasibility of the process although several problemshave to be resolved to allow for a full-scale application,such as the biomass disposal, the purification of theleach liquor before the final manganese recovery, andthe high cost of the process.

3.5. Electro- reductive leaching

Electro-leaching dissolves the manganese minerals inleach solutions whose potential is controlled potentios-

tatically. This method is based on the principle thatmanganese dioxide can be used as an electrode becauseit is a semi-conductor. In the electroslurry process,manganese is cathodically dissolved from oxide slurryby direct electroreduction.

Elsherief (2000) has recently applied electro-reduc-tive leaching for upgrading Egyptian low-grade man-ganese ores in sulfuric acid solutions using a slurryelectrolysis cell. It was found that MnO became readilysoluble in acid under reductive conditions. Cyclicvoltametric studies on carbon paste electrodes of theore and its mineral constituents showed that maximumreduction of manganese occurred at 0 mV with respectto a mercury/mercurous sulfate reference electrode. Theacid concentration, the reaction temperature, and theapplied potential affected the reaction rate and levelof extraction of manganese from the ore. The rate ofreduction increased with increasing acidity. The reac-tions were proposed as follows:

MnO2 þ 4Hþ þ 2e− ¼ Mn2þ þ 2H2O ð12ÞMnO2 þ Hþ þ e− ¼ MnOOH ð13Þ
MnOOH þ 3Hþ þ e− ¼ Mn2þ þ 2H2O ð14Þ
With sufficient accumulation of MnOOH on the sur-face of the MnO2, further reduction to Mn(OH)2 couldoccur:

MnOOH þ H− þ e− ¼ MnðOHÞ2 ð15Þ

The presence of Fe2+ and Mn2+ ions greatly in-creased the speed of reaction. This could be due to aninvolvement of Fe2+ in the chemically reductive reac-tions as shown in reactions (3–5). Temperature wasfound to have a major influence on the leaching process.The main reduction process resulted in Mn2+ releaseand selectivity could be achieved. At 20 °C, very poorrecoveries of metals were recorded. Under optimumconditions at 70 °C in 50 g/L H2SO4 solutions at 0 mV,manganese was fully extracted whereas 56% iron wasextracted in 45 minute electrolysis. The amount ofmanganese leached by the electrolysis at 0 mV wasapproximately five times more than that recovered bychemical leaching without applied potential.

3.6. Simultaneous leaching manganese(IV) oxide andsulfide minerals

Many studies have been carried out to leachmanganeseoxides and sulfide minerals simultaneously in an acidmedium, H2SO4 or HCl. In this leaching process, the

Table 3Simultaneous leaching of Mn oxides and metal sulfide minerals

Sulfides Mn oxides Conditions Key results Ref

Pyrite (FeS2) Low Mn ore H2SO4, 175 °C (1) withO2 (2) without O2

(1) MnSO4.H2O product (2) FeSO4

suitable for leaching at room tempThomas andWhalley 1958

Mn ore 10% H2SO4 Mn ore:pyrite ore=1:1 milled before leaching increased Mn recovery Gaprindashviliet al. 1993

Mn Conc.(58% MnO2)

10% H2SO4 85–95 °C 99% Mn recovery Omarov andBeisembaev 1996

Mn–Ag ores Residue cyanidation 93% Mn recovery 92% Ag recovery Lu and Zou 2001ZnS MnO2 H2SO4 sol. MnSO4, ZnSO4 suitable for electrowinning Li 2000

Mn–Ag ores H2SO4 sol. Ag by NH3 leach 98% Mn recovery 92% Ag (99% pure) Yaozhong 2004FeS2/ZnS/PbS Mn ore H2SO4 sol. Key factor: MnOx:MeS, Temp. Kholmogorov

et al. 1998Ni matte Mn nodules 1.5–2.0 M HCl 95 °C, 80 min N90% Ni and Cu and 80% Co Chen et al. 1992Zn matte Mn nodules HCl solution, 70 °C 97% Mn, 92% Co, and 88% Ni Kuh et al. 2001Pyriti-ferrous lignite Low Mn ore 1.9 M H2SO4 Ore/lignite=

1:1.8 99 °C, 3 hN99% Mn recovery Naik et al. 2002


sulfide minerals function as reductants while the manga-nese oxides as oxidants. The sulfide minerals includegalena (PbS) (Kholmogorov et al., 1998), sphalerite (ZnS)(Kholmogorov et al., 1998; Li, 2000; Yaozhong, 2004) orzinc matt (Li, 2000), pyrite (FeS2) (Thomas and Whalley,1958; Gaprindashvili et al., 1993; Omarov and Beisem-baev, 1996; Kholmogorov et al., 1998; Lu and Zou, 2001),nickel matte (Li, 2000), pyriti-ferrous lignite (Naik et al.,2002). The key operation parameters were found to be theMnOx:MeS ratio, acid concentration, temperature andleach time (Kholmogorov et al., 1998). The leachconditions and major results are summarised in Table 3.

Among the sulfides, pyrite concentrate offers someadvantages:

1. The iron introduced can be easily removed as ironoxides or hydroxides in the presence of oxygen.

2. In the absence of oxygen, Fe(II) can be used as areductant in subsequent leaching process.

Other valuable metal sulfides such as nickel and zinccan only be justified for special cases, e.g. a compre-hensive recovery of these metal values from polyme-tallic manganese nodules.

3.7. Leaching with hydrogen peroxide

Jiang et al. (2003) investigated a simultaneousleaching process for extraction of manganese and silverby one-step leaching in sulfuric acid solution in thepresence of hydrogen peroxide. Thermodynamics of theMn–H2O and Ag–H2O systems show that there is apredominance region where Mn2+ and Ag+ coexist insolution. Hydrogen peroxide plays dual roles in the

process: as an oxidizing agent for native silver and areducing agent for manganese dioxide based on thefollowing reactions:

MnO2 þ H2O2 þ 2Hþ ¼ Mn2þ þ 2H2O þ O2

ΔE∘ ¼ 0:549 Vð16Þ

2Ag þ H2O2 þ 2Hþ ¼ 2Agþ þ 2H2OΔE∘ ¼ 0:971 V

ð17Þ

A recovery of 98% for manganese and 85% for silverwas attained for an ore analysing 12.2%Mn and 1850 g/tAg under the conditions of 0.8 mol/L H2O2, 0.8 mol/LH2SO4, 25 °C, and 2 h of residence time.

3.8. Leaching with hydrochloric acid solution

A hydrochloric acid leaching process for manganeserecovery from manganese ores has been proposed andpatented by Abdrashitov et al. (2001). In this process,the Mn-ore feed was leached with aqueous 5–20% HClsolution, which resulted in the associated evolution ofCl2 and the formation of MnCl2. The leach solutiontogether with MnCl2 and Cl2 was treated in the sub-sequent stage by addition of Na2CO3 to precipitatepurified MnO2 product in the presence of aqueous NaCland evolved CO2. The typical manganese recovery intosolution by the acidic leaching was 97.0% at 101.6 g/LMn, and the MnO2 precipitate in the second stage was98.0% pure compared with 17–50% Mn and 50% SiO2

in the ore feed. The novelty of the invented processutilises the chloride ions as the reductant at high acidityand the Cl2 product as the oxidant to precipitate thereduced manganese as MnO2 in alkaline conditions.However, this process is expected to be rather expensive


and inefficient because a very high acidic condition isneeded for the chloride ions to function as a reductant. Inaddition, the corrosive Cl2 gas has to be collected and tobe used in the subsequent precipitation step.

3.9. Leaching with nitrogen dioxide and nitric acidsolutions

A nitrous-nitric process was developed for recoveryof manganese from manganese ores by Dean et al.(1942) at U. S. Bureau of Mines. The process involvedleaching the ore with NO2 and decomposing the re-sulting Mn(NO3)2 solution. Five hours of nitration ofCuyuna Merritt ore (−65 mesh) gave at least 95%extraction of manganese in a solution low in iron andphosphate. HNO3 formation could be maintained as lowas 0.3 g per gram of manganese by controlling the NO2

flow and eliminating air pressure. Manganese concen-trations of 300 g/L could be obtained. Evaporation ofthe pregnant liquor resulted in quantitative precipitationof Fe and P. Electrolytic decomposition of the filteredsolution, heated by alternative current, proved to be thebest method to decompose the salt and gave a high-purity MnO2. Magnesium nitrates could be removed bybleeding and re-crystallisation.

In a recent development by Drinkard and Woerner(1997), HNO3 has been used in leaching for recoveryof metal values from metallurgical dust wastes. Although

Table 4Reductive leaching of manganese nodules by aqueous SO2/sulfite

Reagents Conditions Key results

(NH4)2SO3 (S1)(NH4)2CO3 (S2)

−100 mesh 80 °C, 120 min50 g/L (S1) 120 g/L (S2)

S1 enhanceCo 98, Cu

SO2–H2SO4 Control the ratio of moleof SO2 to nodule weight

Selective fo

SO2 SO2–NH3 SO2 alone nselective fo

SO2–NH3 (Pressure leach) Pilot plant pH 8–9 (%) 85 Cu,(NH4)2SO3 80–110 °C Ea=62 kJ/m

Fit 1− (1−aSO2(a)–H2SO4

(b)–(NH4)2SO4(c)60 g/L (a), 3% (b), 100 g/L(c) 20% solid, 3 h

(%) 88.5 C97.8 Zn, 99

Na2SO3–(NH4)Cl 10–80 °C 0.5–2.5 h R dependspore diffusichemical R

(1) spent electrolyte(2) 100 g/L FeSO4

(1) 40–80 °C(2) 50 °C, pH 3–4

(1) leach N(2) leach C

FeSO4–H2SO4 90 °C, 30 min N90% Ni, CRoasting NH3/Air leach Roasted at 750 °C

Two stage leaching(1) monitor(2) selectiv

Roasting NH3–CO2 leach 190–200 NH3 85–110 g/LCO2 Leachate aeration

51–57% M

Roasting aq. NH3–(NH4)2CO3 Leach with air Selective (%

nitrous-nitric process is effective, HNO3 is an expensivereagent and the resulting effluent is difficult to treat.

3.10. Acid leaching of Mn(II) carbonate and silicateores and slags

Manganese carbonate and silicate minerals are solu-ble in acids, e.g., in sulfuric and hydrochloric acids.Several acid leaching processes have been developedto recover manganese from these kinds of manganeseminerals.

Arsent'ev et al. (1992) investigated hydrometallur-gical methods for processing of Vorkutinsk manga-nese ores. The ores containing 41.1% Mn consisting of53–55% carbonates and 40–43% silicates were groundand leached with H2SO4 (0.8 kg/kg ore) for 1.5 h atabove 90% Mn recovery. The leach solution was thenneutralized, and a carbonate concentrate was produced,containing 35–45 Mn and 0.2% SiO2 at 98% removal.

Das et al. (1978, 1979) investigated the recovery ofmanganese from ferro-manganese slag by acid leaching.The slag from ferro-manganese manufacture contained12–30% Mn(II) with the main impurities being CaO,SiO2, and Al2O3. The slag was leached with HCl orH2SO4. Manganese recovery was over 90% under opti-mum conditions.

Comba et al. (1991) developed a calcium fluoride-enhanced hydrochloric acid leaching procedure for

Ref

d recovery (%) Ni 93,97 Mn and Feb1

Maslenitskii et al. 1969

r Ni, Co, and Mn over Fe and Cu Pahlman and Khalafalla 1979;Kanungo and Das 1988

on-selective SO2–NH3

r Ni, Co, and CuKawahara et al. 1991

90 Ni, 80 Co 55% Mn in Mn cake Mittal and Sen 2003ol R=k[S(IV)][MnO2]

0.5

)1/3 =ktDas et al. 1998 Demirbas 1999

u, 99.8 Ni, 91.8 Co,.6 Mn and 2.4 Fe

Acharya et al. 1999

on [S1] and [S1] R(Mn, Ni, and Cu)

on R(Fe) surfaceate (Co) jointly controlled

Choi et al. 1996 Li et al. 1997Sventsitskii et al. 2003

i and Cuo and Mn

Kane and Cardwell 1976

u and Mn N85% Co Vu et al. 2005Eh for Co recoverye for Ni and Cu

Jana et al. 1999

n recovery Javorek 1988

) 93 Ni, 92 Cu, 65 Co Saha et al. 1993


recovery of manganese from manganese-bearing silicateores. The proposed process is an alternative to high-temperature and energy-intensive extractive techniques.Single-stage leaching with a stoichiometric amount ofHCl and F-Mn mole ratio of 0.2 extracted 96% Mnwhen the leaching conditions were 50 °C, 3 h, and 35%solids. Manganese at 98% recovery was achieved duringa countercurrent leaching under similar conditions and99% CaF2 was reclaimed.

4. Processing of manganese nodules, heterogeniteand secondary materials

4.1. Processing of manganese nodules

Manganese nodules have been recognized as a po-tential source of various metals such as Cu, Ni, Coand Mn. It is well known that the manganese nodulescontain these metals as higher oxides or hydroxides andusually amorphous in nature. Therefore, it is essential toroast or leach the manganese nodules in the presenceof a reductant. In pre-roasting, non-coking coal can beused as reductant followed by leaching (Javorek, 1988;Saha et al 1993). In direct reductive leaching, variousreductants were investigated including sulfur (Mohantyet al., 1994), SO2 or sulfurous acid (Pahlman andKhalafalla, 1979; Kanungo and Das, 1988), manganousion (Acharya and Das, 1987), glucose (Das et al., 1986),Charcoal (Das et al., 1989) and ferrous iron (Anandet al., 1988). Among these reductants, SO2 and FeSO4

are two of the effective reductants for MnO2 leaching.However, SO2 is preferred for its rapid rate of reac-tion, low temperature operation, ease of purifying leachliquor and elimination of barren solution disposal prob-lem. Importantly, SO2 is cheap to make and can beeasily extracted from spent gas of smelters.

In comparison to manganese ores, the leaching ofmanganese nodules is more complex and difficult due totheir matrix structures and polymetallic value recover-ies. These metals are present in nodules as oxides orhydroxides. The major matrices for Mn and Fe aretodorokite (buserite), δ–MnO2, goethite, maghemite orMn-hematite, where nickel and copper are mainly asso-ciated with manganese oxide phases and cobalt mostlywith the iron phase, but possibly also with the manga-nese oxides. The most stable species of these elementsin the ocean environment are MnO2, Fe2O3, CuO, Ni3O4

and Co2O3 (Vu et al., 2005). The recent developments inprocessing ocean manganese nodules have been re-viewed by Mukherjee et al. (2004). Table 4 summarisestypical studies on reductive leaching of manganesenodules.

4.1.1. Leaching with sulfur dioxideThe higher oxides of Mn Ni and Co in manganese

nodules, if present, would require reductant to convertthe higher oxides to soluble forms. Goethite may giveferric sulfate or is reduced to give ferrous sulfate insolution. The various reactions in the presence of aqueousSO2 have been reported by Acharya et al. (1999):

MnO2 þ H2SO3→MnSO4 þ H2O ð18Þ

Co2O3 þ 2H2SO3→2CoSO4 þ 2H2O ð19Þ

Ni2O3 þ 2H2SO3→2NiSO4 þ 2H2O ð20Þ

2FeOOH þ 3H2SO4→Fe2ðSO4Þ3 þ 4H2O ð21Þ

2FeOOH þ H2SO4 þ H2SO3→2FeSO4 þ 3H2O ð22Þ

The selectivity of Ni, Co and Mn over Cu, Fe and Alwas found to depend on the ratio of the mole of actuallydissolved SO2 in the solution (MSO2) to the weight ofground nodules (Wsn) (Pahlman and Khalafalla, 1979).Optimum ratio of MnSO2/Msn would vary with specificcomposition and amount of nodules. For the IndianOcean nodules which are low in copper, such selectiveleaching may not be economically viable (Kanungo andDas, 1988).

Depending on the reaction conditions, formation ofmanganese dithionate or soluble metal bisulfites maytake place. The Fe(III) present in the solution may pre-cipitate as jarosite in the presence of ammonium sulfateas shown below:

3Fe2ðSO4Þ3 þ 12H2O

þ ðNH4Þ2SO4→2NH4Fe2ðSO4Þ3ðOHÞ6ð23Þ

The rate of leaching with aqueous SO2 was found toincrease with dissolved SO2 concentration up to 80 °C(Demirbas, 1999). In the temperature range of 80–100 °C,the activation energywas determined to be 62–65 kJ/mol,indicating a chemically controlled process (Das et al.,1998; Demirbas, 1999). In sodium sulfite and NH4Clsolution, the dissolution rates of manganese, nickel andcopper were found to be controlled by pore diffusion,whereas iron by surface chemical reactions and cobaltlimited by both surface chemical reactions and pore dif-fusion (Choi et al., 1996).With an increase in temperaturefrom 80 to 100 °C, the ratio of Mn2+/Mn3+ increased byabout five times indicating favourable formation of lowervalency states of manganese ions at higher temperatures(Das et al., 1998). Similar to the leaching of manganeseores, the reaction product of SO2 wasmainly sulfate with a


proportion of dithionate (S2O62−), depending on specific

leach conditions (Popov et al., 1981). The ratio ofdithionate to sulfate was found to increase with increasingSO2 concentration and higher solution pH in the leachsolution.

For processing polymetallic values such as Ni, Co, Cu,Mn, and Fe, the leaching with aqueous SO2 alone wasfound to be non-selective (Kawahara et al., 1991; Li et al.,1997). A selective extraction of Ni, Co, and Cu from Mnnodules can be achieved by using ammonia/carbonatesolution in the presence of sulfur dioxide as the reductant.Strategies in the presence and absence of O2 (air) can beused to selectively complex metal values in the solution orto precipitate basic manganese carbonate via heating andrecovery of ammonia (Rokukawa 1990). Pre-treatment ofthe ores with dilute acetic solution was found to beeffective to remove a substantial quantity of calcareousand carbonaceousmaterial from the nodules (Pahlman andKhalafalla, 1988). An increase in temperature suppressedthe dissolution of iron into the leach solution. The leachingresidue was mainly composed of manganese ammoniumsulfite hydroxide ((NH4)2Mn2(SO3)3) in an ammoniasolution while manganese carbonate was detected in theleaching with ammonium carbonate solution. Most of Cu,Ni, and Co was extracted into solutions, leaving mostmanganese in the residue at 80 °C for 120 min.

A polymetallic manganese nodules pilot plant of500 kg per day processing capacity was designed and setup at Hindustan Zinc Limited, Udaipur (India) (Mittaland Sen, 2003). The process was based on the reductivealkaline pressure leaching of nodules using sulfur diox-ide as reducing agent in the presence of ammonia at pH8–9. The overall process scheme is shown in Fig. 2.

Fig. 2. Schematic flowsheet of manganese nodule pilot operation us

In this process, the partially dissolved manganeseduring leaching along with copper, nickel and cobaltwas precipitated and separated out as manganese diox-ide in a demanganisation autoclave by sparging oxygen.Free ammonia present in the leach liquor was recoveredin an ammonia stripper and recycled back to leachautoclave as recycled ammonia. Ammonia free leachliquor was subjected to single stage copper extractionwith LIX 84I at controlled pH for selective removal ofcopper. The copper raffinate containing dissolved nickeland cobalt was subjected to sulfide precipitation usingsodium sulfide. The nickel and cobalt were precipitatedas sulfides along with minor impurities like copper, zinc,iron etc. The cake was dissolved in a dilute sulfuric acidsolution in the presence of oxygen under medium tem-perature and pressure conditions. The cobalt and nickelin the purified leach solution were extracted with PC-88A and D2EHPA respectively and produced as puremetal through electro-winning. The experiments con-ducted in the pilot plant showed an average recoveryfor copper, nickel and cobalt as 85%, 90% and 80%,respectively.

The process has several advantages and compareswell with other alternative hydrometallurgical processesfor extraction of metals from poly-metallic nodules.First, the major reagent is sulfur dioxide, which is com-paratively cheap and can also be economically extractedfrom smelter off-gases. Second, the high recoveriesfor nickel, copper and cobalt values can be achieved,which not only covers themining cost but gain additionalprofits. Last, the manganese cake containing 55% man-ganese can directly be used for the production of high-grade ferro-manganese.

ing SO2 reductant in Indian (based on Mittal and Sen 2003).


The National Metallurgical Laboratory (NML) ofIndia developed a two stage ammoniacal leachingscheme with a prior pre-conditioning step (Jana et al.,1999). The necessity of monitoring redox potential in thefirst stage leaching to control cobalt recovery has beenemphasized. Based on the two stage leaching scheme,recycle leaching has been carried out to generate leachliquor having suitable composition for the subsequentsolvent extraction-electrowinning operation. The aver-age recovery of metals in 16 cycles of leaching has beenfound to be 92% Cu, 90% Ni and 56% Co.

4.1.2. Leaching with FeSO4

A two-stage leaching process for selectively extract-ing metal values from manganese nodule was developedby Kane and Cardwell (1976). In the first stage leaching,spent electrolytes containing about 4% free acid wereused to leach the nodule for 14 h at 40–80 °C forpreferentially extracting nickel and copper values. Inthe second stage, the filtered solids were reductivelyleached with solutions containing approximately 100 g/LFeSO4 at about 50 °C and pH 3–4 to extract cobaltand manganese. Air was bubbled through the leachsolution in the second stage to remove iron as oxides/hydroxides to the residue. Copper was extracted at pH 2from the first stage leach solution using a α-hydroxyoxime or 8-hydroxyquinoline. The copper strip solutionwas electrolysed to give the cathodic copper product,and the raffinate from the solvent extraction was simi-larly treated in a nickel cell for the recovery of cathodicnickel. After iron was removed from the second leachsolution, cobalt was extracted by the organic solution atpH 3.5–4.5, and the raffinate was electrolysed for therecovery of manganese.

Leaching of manganese nodule in ammoniacal me-dium using FeSO4 as the reductant was investigated byAnand et al. (1988). Ferrous iron in ammonia-ammo-nium solution can form a sable ferrous ammine complexin the pH range of 9.5 to 9.8. It is oxidised and pre-cipitated as ferric hydroxide by simultaneous reductionof MnO2, Ni3O4, Co2O3 minerals to lower-valencemetal ions which can be stabilised in the solution asmetal ammines. While maintaining the total molarity ofammonia and ammonium sulfate at 5.85 M, most of thecopper, nickel and cobalt, and about 32% of manganesecould be leached within 10 min. An extension of leachtime resulted in partial precipitation of manganous am-mine as MnO or Mn2O3 and the loss of nickel and cobaltdue to coprecipitation or adsorption. This process couldeliminate iron in the solution in the ppm level. However,it was difficult to optimise the selectivity of Ni, Co andCu over manganese by varying the total concentration of

ammonia and ammonium sulfate in the range of 1.96 to5.85 M. Other variable affecting the extent of extractionof Ni, Co, Cu and Mn included the amount of ferroussulfate, leaching temperature and initial pH.

More recently, the leaching of manganese nodulewith FeSO4–H2SO4–H2O solutions has been reported(Vu et al., 2005). The manganese deep ocean nodulesoriginated from the IOM area, located in the Clarion–Clipperton ore field. The optimum conditions estab-lished were as follows: stoichiometric amount of FeSO4,1.6-fold excess over stoichiometric amount of H2SO4,90 °C, l/s of 7:1, grain size ≤1000 μm. Under theseconditions, more than 85% of Co and 90% of Ni, Cu andMn could be extracted within 30 min. The leach liquorscontained approximately 146 mg/L Co, 1.63 g Ni/L,1.69 g Cu/L and 30 g/L Mn.

Compared with a non-selective leaching process,the two-stage leach process enable nickel to be sepa-rated from cobalt at early leaching stages and makethe subsequent separation process simple and easy, but theenergy consumption is expected to be high in the leachingstages. The applicability of the process would largelydepend on the valuable metal grades in the nodules andthe availability of cheap spent industrial pickle liquors.

4.1.3. Leaching with nitrogen dioxide and nitric acidNitrous-nitric method was tested for processing

ocean nodules to recover the manganese value selec-tively (Welsh and Sochol, 1978). The nodules com-prised primarily oxides of manganese and iron withlesser amounts of Cu, Ni, Co, alkali metal and alkalineearth metal compounds. The process was comprised ofleaching the nodules at 80 °C with an aqueous solutionof nitrous and nitric acids; removing insoluble ironoxide and gangue at pH b2.5 by filtration, adding Mn(OH)2 to the filtrate to selectively precipitate Cu, Ni andCo metals as hydroxides and adding to the filtrate H2S toprecipitate metal traces as sulfides to obtain purifiedmanganese sulfate solution.

4.1.4. Leaching with cuprous copper solutionSzabo (1976) developed an ammoniacal cuprous

leach process for recovery of metal values from man-ganese deep sea nodules. This process has been tried at apilot plant. The process consisted of ore preparation,reduction-leach, oxidation and wash-leach, LIX separa-tion of the metals, and electrowinning. The noduleswere leached with an ammoniacal solution containingCu(I) ions to reduce the manganese oxides and extractCu, Ni, Co and Mo into solution. The cuprous ions usedfor the leaching were regenerated by reduction of the Cu(II) with a reducing gas such as a mixture of CO 95%


and H2 5%. The nodule residue was washed with anammoniacal-(NH4)2CO3 solution to remove entrainedmetal values and the wash effluent contained Cu, Ni,Co, and Mo values. Rappas (1985) reported a processfor the extraction of metals from laterite ores and oceanmanganese nodules with a stabilised acidic cuprous ionsolution (Cu+L) in a suitable reactor at a temperature upto 90 °C and pH in the range of 1.5–2.5, wherein L is astabilising ligand selected from the group consisting ofCO, XRCN and Cl−, X being H or OH and R beingaliphatic compound having 1–4 carbon atoms.

The cuprous copper process is considered less suit-able for practical applications due to the unstable natureof the cuprous ions in the leach solutions. Sufficientligands or complexing agents are needed to present inthe leach solution and also in the relatively expensive re-generation process.

4.2. Treatment of heterogenite

In recent years a newmanganese bearing ore has drawnattention. It is heterogenite from Africa, containing cobalt,cupper and significant quantities of manganese. Afrowsheet was developed and piloted at Mintek during1998 for recovering Cu and Co from the Kakanda(Democratic Republic of Congo) tailings (Dry et al.,1998; Sole et al., 2005). This was an acid/SO2 reductiveleaching process. Copper in the leach solution wasrecovered by SX-EW. A bleed of the raffinatewent through Fe and Al removal by neutralisation,manganese (about 1 g/L) recovery or removal by solventextraction using D2EHPA, and cobalt recovery by solventextraction using Cyanex 272 followed by electrowinning.

Recently, processing of African Heterogenite atRubamin Ltd, India has been reported by Das (2006).A similar reductive acid leaching with SO2 as reductantis used under controlled pH. The leach liquor containingMn, Zn, Cu, Ni, and Fe along with Co is first neutralisedwith lime to precipitate most of the Fe and Cu, followedby extraction of Mn and Zn from the leach liquor byD2EHPA, and then extraction of cobalt by phosphonicacid (P507, produced in China). Nickel in the raffinate isrecovered as nickel carbonate. The electrolytic cobaltproduced at Rubamin has a purity of 99.9%. The overallrecovery of cobalt from heterogenite is more than 90%together with production of a wide range of salts ofcobalt, copper, nickel, and manganese.

4.3. Treatment of waste batteries

Lindermann et al. (1994) described the Batenus pro-cess for recycling battery waste. The crushed batteries

were leached in sulfuric acid. Several techniques wereemployed for the separation of metals.

1. The insoluble manganese dioxide and carbon residuewere sold to the ferro-manganese industry and themetal salt solution was treated by ion exchange torecover copper, nickel, cadmium, and mercury. Thezinc was recovered by solvent extraction.

2. The manganese from the leach solution was pre-cipitated as carbonate from the raffinate.

3. The remaining alkaline metal sulfate and chloridesolution were separated into acid and base by electro-dialysis with bipolar membranes.

The Batenus process was reported to be free of anyemission, and produce raw materials with remarkablyhigh purity. The zinc, nickel, cadmium, copper, andmercury were separated out of the used zinc-carbon,alkaline manganese, and nickel-cadmium batteries.Moreover, the manganese carbonate, obtained as rawmaterial, could be used for the production of manganesedioxide for new batteries.

A hydrometallurgical process for the recovery of zincand manganese from spent batteries was developedby Bartolozzi et al. (1994). In this process, the spentbatteries was leached with 32% sulfuric acid, and afterliquid and solid separation, the dissolved iron wasremoved by precipitation by NaOH at pH 3.7 and thenby 32% ammonia at pH5 and 60 °Cwith addition of smallamount of H2O2. The subsequent solution was firstelectrolysed to remove Hg, Ni, and Cd, and then at pH 3.7was electrolysed to recover cathodic zinc and anodicMnO2. TheMn(IV) in the leach residue was solubilised inacidic hydrogen peroxide (35%) and re-precipitated withNaOH at 40 °C. This process appears feasible technically,but hardly likely to be applied economically with sostrong conditions and the use of expensive reagents suchas H2O2 and NaOH to recover manganese dioxide.

For treatment of waste lithium battery, Shibata andBaba (1995) investigated a process in which lithium andmanganese were selectively leached from the battery byusing diluted mineral acids HCl and HNO3. Lithium andmanganese were precipitated as their carbonates fromthe individual leached solution with sodium carbonatesolution.

A process for recovery of metal values from chemicalby-products of spent Mn–Zn batteries was reportedby Ptitsyn et al. (2001b). The by-products from spentMn–Zn batteries were processed by grinding, magneticseparation of iron scraps, firing the powders, acidicleaching, and electrowinning of zinc on the cathode andMnO2 on the anode.


4.4. Treatment of Mn-bearing sludges and slags

Kukula and Sikora (1977) reported a hydrometallur-gical process for the recovery of manganese fromslag leaching solutions. The solution contained (g/L)SiO2 15–20, Mn 25 and (NH4)2SO4 197, and (mg/L) Cr1.5, Ni 2.76, Co 7.78, Cu 5.4, Mo 0.95 and Fe 800. Thepurificationwas accomplished by the followingprocedure:

1. SiO2 was removed by precipitation with lime.2. Fe2+ was precipitated as Fe(OH)3 after oxidation with

KMnO4.3. Nickel and cobalt were removed partly by co-precip-

itation with iron, partly by reduction with manganesedust added.

4. The excess of cobalt was extracted by solvent extrac-tion, and the organic impurities were removed byadsorption on active carbon.

5. From the pure solution, manganese was recoveredelectrolytically on aluminium cathode at current den-sity 300–400 A/m2 in a membrane system.

The feasibility of this process would depend on theinitial concentrations of manganese and other valuablemetals such as cobalt and nickel in the solution. For thisparticular case, the initial cobalt content is consideredtoo low to be feasible.

A process for the recovery of manganese fromelectrolytic zinc anodic slime and from scrap dry cellsdescribed by Yoon and Kim (1986) involved the fol-lowing steps: (a) manganese was extracted from anodicslimes of electrolytic zinc anodic slime and scrap drycells by leaching with HCl. (b) The impurities (Zn, Fe,Pb) were removed from the leach solution prior toelectrolysis by neutralization with NH4OH and cemen-

Fig. 3. Conventional process for production of electrolytic man

tation with manganese powder. (c) The manganese waselectrodeposited from the purified chloride solutionwith added Na2SeO4. However, evolution of Cl2 duringleaching and electrowinning are likely to be issues andconsumption of ammonia is expected to be high.

A process for recovery of Zn and Mn from a watertreatment plant sludge was developed (Pesic et al., 1996;Binsfield et al., 1996). The process features leaching ofthe sludge with H2SO4 in the presence of SO2 reductantat pH 2 and 60 °C for 15 min, three-stage purification,and simultaneous electrowinning of zinc and MnO2.

5. Process for recovery of manganese from solution

5.1. Manganese electrowinning-production of electrolyticmanganese

Electrowinning of manganese is a major process forproduction of high purity electrolytic manganese (EM).The electrowinning systems can be classified based onthe media such as sulfate and chloride because all theparameters differ substantially in different media.

Electrowinning of manganese in sulfate solutions hasbeen conventionally employed for production of mostmanganese metal required mainly for steel alloys, par-ticularly in special steels such as low carbon stainlesssteel grades. The overall process for production ofelectrolytic manganese from ores consists of a series ofsteps as presented in Fig. 3.

Iron(III) is typically removed by hydroxide precip-itation at pH 3–4 and other impurities such as Ni,Co, Cu, Zn are removed by sulfide precipitation. Theresultant MnSO4 solution can be used for productionof electrolytic manganese. This process was firstdeveloped by Fuller et al. (1964) and was implemented

ganese (EM) and electrolytic manganese dioxide (EMD).

Table 5Electrolysis of manganese from sulfate solution (Pisarczyk 1995)

Condition Value

Feed solution, catholyteMnSO4 30−40, g/L(NH4)2SO4 125−150, g/LSO2 0.3−0.5, g/L

AnolyteMnSO4, g/L 10−20, g/LH2SO4, g/L 25−40, g/L(NH4)2SO4, g/L 125−150, g/LCurrent density, mA/cm2 43–65Catholyte pH 6–7.2Anode composition Pb+1% AgCathode composition Hastelloy, type 316Cell voltage 5.1 VTemperature 20 °CDiaphragm AcrylicCurrent efficiency 60–70%


by Harris et al. (1977) in South Africa. The plant oper-ated at its rated capacity of 16,500 t/year at currentefficiencies of approximately 68%. The typical con-ditions required for the electrolysis process are sum-marised in Table 5 (Pisarczyk, 1995).

The concentration of MnSO4 was found to be one ofthe key parameters required for achieving a high currentefficiency (Jakuszewski et al., 1972). In electrolysis,both anodic and cathodic yields were improved withan increased MnSO4 concentration. An increased(NH4)2SO4 concentration gave a higher yield of cathodicmanganese with an improved quality, while an increasedH2SO4 concentration gave a smaller yield of relativelypure MnO2 at anodes (Jakuszewski et al., 1972).

A number of studies on electrowinning of manganesein chloride electrolytes have been reported in the liter-ature (Jacobs and Churchward, 1948). Compared withthe sulfate system, the electrolytic process in chloridemedium offers some potential advantages and disad-vantages (Agladze and Muchaidze, 1957a):

1. Lower cell potential can be used, resulting in up to25% less power consumption. Generally, the efficien-cies are higher than the sulfate systems by 5–10%.

2. 2–3 times higher current density can be achieved(Parissis and Winand, 1992), thereby decreasing thesize of the electrolysis plant.

However, it has potential disadvantages:

1. Large amounts of ammonia in the electrolyte maydecompose at the anode at approximately 0.3 t perton of manganese.

2. Evolution of chlorine at the anode which has to beremoved or destroyed.

These disadvantages have limited the application ofmanganese electrowinning in chloride medium.

5.1.1. Effect of additivesAn addition of 0.15–5.0 mg/l of Zn2+ to the catholyte

in the manganese deposition cell has been reportedto increase the cathode current efficiency (Mantell andFerment, 1969; Ilea et al., 1997) and decrease theminimum current density necessary for manganese de-position on the cathodes. This occurred in both puresolutions and solutions containing trace amounts ofimpurities in manganese ores, e.g., Co, Ni, Mo, and Cu.This effect was observed for the catholyte containing(g/L) Mn2+ 35, (NH4)2SO4 120, and SO2 0.1 at 30 °Cand pH 7.2, with the anolyte containing (g/L) MnSO4

15, H2SO4 30, and (NH4)2SO4 130. However, in thepresence of Zn2+ greater than 5.0 mg/L, the purity of thedeposited manganese was decreased to below 99.9%.

The presence of selenium and tellurium ions in theelectrolyte was also found to be beneficial for the cath-odic current efficiency and specific energy consumption(Louis and Martin, 1976; Ilea et al., 1997). For anindustrial electrolyte containing 105 g/L MnSO4 and1 M (NH4)2SO4 in the presence of 1 mM H2SeO3, thecurrent efficiency reached 82% and specific energyconsumption was 6.8 kWh/kg for a conversion of about44%, at 500 A/m2, pH 7–8 and 20 °C (Ilea et al., 1997).Using Pt or Al cathode and a Pb–Ag anode, seleniumwas found in the anode mud, as higher valent manga-nese selenites (Suliakas et al., 1973). Minimum amountof the anodic mud was produced at pH 2 and 40 °C at100 g/l H2SeO3. When selenium was present as SeO4

- ,an insoluble selenium compound appeared only at tem-perature greater than 70 °C and was independent of theelectrolyte pH.

A natural or synthetic organic H2O-soluble polymer,e.g., a H2O-soluble polyacrylamide or natural guargum in combination, led to less mechanical work inremoving the deposited manganese from the cathodesand an increase in both the quantity and quality of thedeposit (Coleman and Griffin, 1984). An addition ofthiourea in the range 0.015–0.03 g/L was found topromote current efficiencies up to 60–68% (Hammer-quist, 1951).

5.1.2. Impurity controlSince manganese has a high negative value in the

electromotive series, the levels of metals with higherreduction potentials have to be controlled to a low level.

Table 6The maximum admissible impurity content in industrial manganeseelectrolytes

Elements Maximum admissible (mg/L) References

Al 1000 Ilea et al. 1997Co 0.5Fe 10Cu 5Ni 0.1–1 O'Keefe 1984Zn 20 Ilea et al. 1997

N5 affect Mn quality Mantell and Ferment 1969Mo 1.5 Anon. 1949


Typical impurity contents in the industrial electrolyteand their admissible concentrations are given in Table 6(Ilea et al., 1997). Not all the metals in the range affectthe polarization (Mantell, 1944). Ni, Co, Ag, Cu, Zn andMo had some effects while Cd, Mg, and Na was foundto have no effect.

In a patented method (Mantell and Hammerquist,1944), the level of magnesium in the anolyte was con-trolled by fluoride precipitation. For the anolyte solutioncontainingMnSO4 1–15 g/L and (NH4)2SO4 150–75 g/L,magnesium in the anolyte was precipitated as MgF2by addition of a soluble fluoride, e.g., NH4HF2 in acontrolled pH range of 1.5–2.75. The precipitate was thenremoved by filtration. The concentration of magnesiumleft in the solution was below 2 g/L. The filtrate waspassed to the leaching tank and used for leaching ofmanganese ores.

Nickel, cobalt, and copper are the common impuri-ties in the electrowinning of manganese. The electro-plated manganese can couple the electroplated Ni, Co,and Cu to cause redox reactions resulting in hydrogenevolution. A study on the corrosion rate in a solutioncontaining 25 g/L Mn and 150 g/L (NH4)2SO4 at pH 7and 20 °C showed that the Mn–Co and Mn–Ni coupleshad slightly greater corrosion rates than manganesealone with relative influence in the order CoNNiNCu(Agladze and Gofman, 1957).

The presence of molybdenum seriously affects theefficiency of production of high-quality EMD. Theupper critical maximum concentration of molybdenumwas established to be 1.5 mg/L (Anon, 1949). A proce-dure was developed for the control of Fe, Mo, Ni and Coin the electrolyte (Wanamaker and Morgan, 1943; Anon,1949; Wanamaker et al., 1951). It consisted of threemajor steps:

1. The spent electrolyte was oxidised with the cell sludgeas well as additional manganese ore in which man-ganese must be present in a valence greater than 2.

2. After digestion, the pH was adjusted to 5.5–6.0 toprecipitate Fe and some of Mo as sludge.

3. The pH was then adjusted to 7.0–7.5 and (NH4)2Sadded to precipitate the Mo, Co, and Ni.

The presence of silica is detrimental to the formationof a satisfactory cathodic deposition in manganese dia-phragm electrowinning. A coagulation method with Al2(SO4)3 as the coagulant was developed (Hammerquist,1948). The process required the solution pH in the rangeof 5.5–7.5 and the ratio of 20–25 kg of Al2(SO4)3 toabout 100,000 kg of electrolyte. The precipitate, Al(OH)3, could then be filtered off.

5.1.3. Effect of anode typesThe type of anode plays an important part in man-

ganese electrowinning. In addition to the physicalproperties such as resisting breaking and bending, theelectrochemical properties in terms of resistance tocorrosion and allowance for low potential for O2 evo-lution are major concerns. Lead or lead alloy anodesare conventionally used in manganese electrowin-ning in sulfate medium while in chloride medium, Cor graphite is usually employed. Various new types ofanodes containing composite elements have been de-veloped. The novel anode types include an alloy of Pb,Sn, Sb and Co (Fink and Kolodney, 1939), a compositeanode containing 0.5–5% Ni and 1–10% Mn, andbalance Ti alloy (Debrodt et al., 1986), a compositeanode consisting of 0.50–2.4% Ag, 0.25–2.5% As, andthe balance Pb (Godsey, 1955), an anode of Ti, Zr, orTa coated with a layer of MnO2 (Osawa, 1973), and Tianodes (Ptitsyn et al., 2001a). These composite alloysgenerally improve the life of the anode and lowerthe cell energy consumption. However, the compositeanodes are costly and the application of these noveltypes of anodes largely depend on the balance of costand benefits offered.

5.2. Production of electrolytic manganese dioxide (EMD)

EMD can be made by the anodic oxidation ofMnSO4 solution with inert electrodes. The cell reactionsare:

Anode : Mn2þ þ 2H2O→MnO2 þ 4Hþ þ 2e− ð24Þ

Cathode : 2Hþ þ 2e−→H2 ð25Þ

Overall : Mn2þ þ 2H2O→MnO2 þ 2Hþ þ H2 ð26Þ


The mechanism of deposition of MnO2 on anodewas proposed to form initial Mn3+ which disproportionsaccording to:

Mn3þ þ 2H2O→MnO2 þMn2þ þ 4Hþ ð27Þ

A typical process for production of EMD fromMnO2

or MnCO3 ore includes reductive roasting if MnO2

ore is used followed by sulfuric leaching and purifica-tion steps (Taylor, 1993) which are similar to those forpreparation of EM solutions (see Section 5.1).

The removal of potassium is required for productionof electrolytic manganese dioxide because the presenceof potassium during electrolysis promotes the formationof the α-MnO2 which is non-battery active. This isusually achieved by jarosite or alunite precipitation inthe leaching process in which ferric sulfate is normallyadded (Paixao et al., 1995). A jarosite precipitationreaction can be represented by:

Kþ þ 3Fe3þ þ 2SO2−4 þ 6H2O

¼ KFe3ðSO4Þ2ðOHÞ6↓ þ 6Hþ ð28Þ

BHPMinerals International built and commercialisedan EMD plant comprising conventional reductive roast-ing, leaching, purification steps in Newcaslte, Australiausing a high manganese ore in 1991 (Taylor, 1993).

5.3. Production of chemical manganese dioxide (CMD)

Chemical manganese dioxide (CMD) can be pre-pared by various methods including thermal decompo-sition of manganese salts such as MnCO3 or Mn(NO3)2under oxidising conditions. CMD can also result fromthe reduction of higher valent manganese compounds,eg, those containing the MnO4

− ions.The conventional process for production of CMD is

via a carbonate route (Raghavan and Upadhyay, 1999).In this process, manganese dioxide ore was treated byreductive roasting followed by sulfuric leaching. Afterremoval of impurities by neutralisation to the pH rangeof 4–6 and filtration, manganese in the solution could besent to electrolysis to produce electrolytic manganesedioxide (EMD) or recovered as MnCO3 by addition of(NH4)2CO3. MnO2 (CMD) was obtained by heating theMnCO3 in the presence of oxygen or air at about 500 °Cbased on the reaction:

MnCO3 þ 1=2O2→MnO2 þ CO2 ð29Þ

The manganese dioxide and manganese carbonatemixture was further leached with sulfuric acid to removed

unroasted manganese carbonate and sodium and potassi-um salts followed by washing and drying to yield theCMD product with 90% as MnO2. A fully hydrometal-lurgical process for production of CMD was proposedby Raghavan and Upadhyay (1999), which used Na2SO3

reductive leaching instead of conventional reductiveroasting.

A commercial process for recovery of manganesedioxide is based on thermal decomposition of man-ganese nitrate solution to produce a substantially pureCMD (N99.5%) (Welsh, 1973, 1978). The processinvolved heating a crude Mn(NO3)2 solution fromleaching Mn-bearing ores at 90–100 °C and adjustingpH to 4.5–5.0 with addition of MnO obtained by thereduction of MnO2-containing ores. The solution wasfiltered and continuously added to a stirred reactor heldat 135–146 °C, where the Mn(NO3)2 decomposed toMnO2 and NO2 according to the reaction:

MnðNO3Þ2→MnO2 þ 2NO2 ð30Þ

The NO2 generated could be recycled by reactionwith water:

NO2 þ H2O→HNO2 þ HNO3 ð31Þ

The acid formed then reacted with low grade manga-nese dioxide ore forming slurry of the crude manganesenitrate to feed the process.

Recently, Pagnanelli et al. (2007) investigated threechemical procedures for preparation of chemical man-ganese dioxide (CMD) for possible applications in drycell batteries instead of electrolytic manganese dioxide(EMD). The three preparation procedures were:

(a) precipitation–oxidation by air plus acid activation(two-step-air)

(b) precipitation–oxidation by H2O2 plus acid acti-vation (two-step-H2O2), and

(c) precipitation–oxidation by KClO3 (single-step-ClO3)

The two-step procedures gave larger throughputs ofactivated manganese dioxide (80–86%) compared withsingle-step preparation. The effect of operating condi-tions on the chemical, structural and electrochemicalproperties of CMDs produced under optimised condi-tions was investigated and compared with those of acommercial EMD sample. The two CMD samples pre-pared using the two-step procedures demonstrated, bypotentiometric titrations, the same acid–base properties

Fig. 4. A schematic diagram for low-dithionate acid leach process (based on Ward et al., 2004; Ward, 2005).


which had the structure and proton insertion propertiessimilar to the commercial EMD, but different structuraldefects. X-ray and IR spectra showed highly disorderedsystems of the CMD samples and the presence of Mn2O3

in first step products and of γ–MnO2 in second stepproduct. Voltammetric cycles denoted that CMD sam-ples obtained after acid digestion present similar peaks tothe commercial EMD but with a higher current intensity.These are encouraging results for further investigationsof CMD and its characterisations in comparison to EMDfor applications in battery industry.

5.4. Recent hydrometallurgical processes for productionof CMD and EMD

Chemical manganese dioxide is conventionally man-ufactured by a roasting–leaching process as shownearlier. However, this is economically justified onlyfor the manganese ores containing N40% Mn. Somepure hydrometallurgical processes for production ofCMD have been developed in recent years. One ofthe processes has been proposed by Raghavan and

Upadhyay (1999) for recovery of manganese as CMD orEMD from manganese ores and other manganese bear-ing materials such as anode mud. It involves aqueousSO2 and H2SO4 leach at pH 2.5 and 90 °C, purificationwith limestone to remove the iron present, crystal-lisation at 4 °C to control the build up of sodium sulfate,precipitation of manganese with addition of sodiumcarbonate at 70 °C and calcination at 500 °C. Thepurification of the calcine to remove unroasted manga-nese carbonate could be achieved by leaching with 5–10 g/L H2SO4 to produce quality MnO2.

In processing of manganese nodule ores with SO2 asreductant, one of the problems is the generation of aproportion of dithionate along with sulfate productduring the reduction. The presence of dithionate wouldresult in a slow release of SO2 from the MnSO4 product.

A low-dithionate acid leach process for the treatment oflow-grade manganese ores and Mn-waste sourcesfor manufacture of high-quality EMD for the alkalinebatterymarket has been developed byHiTech Energy Ltd.(Ward et al., 2004; Ward, 2005). The process featuresleaching with acidic solution at b60 °C and controlled


potential (Eh) for addition of SO2 gas and manganeseore feed to minimise generation of dithionate ion. Themanganese in the leach solution was extracted using acarboxylic acid, e.g. Versatic 10, at pH 5.5–6.5 followedby stripping with spent aqueous H2SO4 electrolyte toobtainMnSO4 feed solution for electrowinning to produceEMD (MnO2). A schematic diagram of the improvedProcess II vs conventional Process I is shown in Fig. 4.

6. Conclusions

The major and potential manganese sources, routesand products are graphically summarised in Fig. 5.

Rich manganese ores (N40% Mn) are dominantlyprocessed by pyrometallurgical processes to produceferro-Mn and silico-Mn alloys, which are mainly used iniron and steel industry. Production of CMD, EM andEMD are mainly accomplished by reductive roastingfollowed by hydrometallurgical processing. Low man-ganese ores (b40% Mn) and polymetallic manganesenodules currently cannot be economically processed bypyro-processes because too much energy is required.

A number of hydrometallurgical processes have beendeveloped, both in acidic and basic media. Amongthe various methods developed, the reductive leaching ofMn(IV) to Mn(II) has been extensively investigated,including the reductants of aqueous SO2, ferrous salts,

Fig. 5. A schematic diagram of major and potential m

reducing organics, H2O2, sulfide minerals, bio- andelectro-reduction. Among these processes, the sulfurdioxide reductive leaching process has been developed toa pilot scale for processing of manganese ores andmanganese nodules. One problem to use SO2 as thereductant is the formation of dithionate by-product. It canbeminimised by controlled the addition of SO2 to achievea desired solution Eh as developed byHiTech Energy Ltd.

However, the crucial issue is the purification ofleaching liquors due to the extraction of iron and man-ganese in the course of leaching operations. Selectiverecovery of Cu, Ni and Co is often difficult from solutionscontaining soluble iron and manganese as contaminants.The leaching processes of ocean manganese nodules canbe conducted in near ambient conditions, using cheap andreadily available chemicals. Selective recovery of Cu, Zn,Ni and Co components is possible over manganese withammoniacal sulfite alkaline leach. In this strategy, Cu, Zn,Ni and Co metals are extracted into solution as ammineswhilst iron and manganese is stabilised in residue ashydroxide, oxides and carbonate.

The main methods for recovering manganese from theleach solution include electrolysis for EMD production,electrowinning for EM production and chemical processfor CMD production. Purification steps are needed,depending on what product is produced. Generally, ironand aluminium impurities are removed by hydroxide

anganese sources, process routes and products.


precipitation and other base metals removed by sulfideprecipitation. Jarosite precipitation is required for removalof potassium ions for the production of EMD. In recentdevelopment by HiTech, solvent extraction has been usedto recover manganese from the leach solution and enableearly rejection of the majority of impurities.

One of the potential sources of manganese, whichhas not received much attention, is the recovery ofmanganese from industrial waste solutions, particularlyfrom Ni-Co laterite high pressure acid leach effluentswhich contains about 1–5 g/L Mn together with otherimpurities. This is the focus of Part II of this reviewwhich considers the application of various solvent ex-traction reagents and precipitation methods.

Acknowledgments

The authors would like to thank CSIRO Mineralslibrary staff for collecting some reference papers, and DrDavid Muir for reviewing this paper and providingvaluable comments.

References

Abbas, H., Askar, M.A., Abd-Elaziz, E.M., 1999. Recycling of zinc–carbon spent batteries: I. Production of manganese and zinc assulfates. Egyptian Journal of Chemistry 42 (4), 361–373.

Abbruzzese, C., 1987. Aqueous SO2 processing of manganese ores. In:Davies, G.A. (Ed.), Separation Processes in Hydrometallurgy,Society of Chemical Industry. Ellis Horwood Limited, London,pp. 77–87.

Abbruzzese, C., 1990. Percolation leaching of manganese ore byaqueous sulfur dioxide. Hydrometallurgy 25 (1), 85–97.

Abbruzzese, C., Duarte, M.Y., Paponetti, B., Toro, L., 1990. Biologicaland chemical processing of low-grade manganese ores. MineralsEngineering 3 (3–4), 307–318.

Abdrashitov, Y.M., Dmitriev, Y.K., Zakharova, N.V., Lakeev, S.N.,Lapin, V.V., Loginov, O.N., Pavlov, A.I., Silishchev, N.N., Sorokin,I.A., Shapovalov, V.D., 2001. Acidic leaching for Mn recoveryfrom manganese ores with oxide product precipitation. RU PatentNo. 2176679.

Acharya, C., Kar, R.N., Sukla, L.B., 2001. Microbial extraction ofmanganese from low grade manganese ore. Transactions of theIndian Institute of Metals 54 (3), 99–103.

Acharya, C., Kar, R.N., Sukla, L.B., 2003. Studies on reaction mechanismof bioleaching of manganese ore. Minerals Engineering 16 (10),1027–1030.

Acharya, R., Ghosh, M.K., Anand, S., Das, R.P., 1999. Leaching ofmetals from IndianOceannodules in SO2–H2O–H2SO4–(NH4)2SO4

medium. Hydrometallurg 53 (2), 169–175.Acharya, S., Das, R.P., 1987. Kinetics and mechanism of reductive

ammonia leaching of ocean nodules by manganous ion. Hydro-metallurgy 19, 169–186.

Agladze, R.I., Gofman, N.T., 1957. The corrosion of manga-nese containing the impurities nickel, cobalt, and copper in themetal. Elektrokhim. Margantsa, Akad. Nauk Gruzin. S.S.R. 1,15–24.

Agladze, R.I., Muchaidze, N.N., 1957a. The recovery of metallicmanganese by the electrolysis of chloride solutions. Elektrokhim.Margantsa, Akad. Nauk Gruzin. S.S.R. 1, 341–353.

Ali, E.A., Shahin, M.A., Ahmed, M.S., Ibrahim, I.A., 2002. Leachingof low-grade Sinai manganese ore using lactose as a reductant.Egyptian Journal of Chemistry 45 (2), 359–374.

Anon., 1949. Electrolyte purification in manganese electrowinning.GB Patent No. 627208.

Anon., 2003. Date accessed: 15/6/2006 The Electrolytic ManganeseDioxide Market. http://www.hitec-energy.com.au.

Anand, S., Das, S.C., Das, R.P., Jena, P.K., 1988. Leaching of man-ganese nodule in ammoniacal medium using ferrous sulfate as thereductant. Metallurgical Transactions B 19B, 331–335.

Arsent'ev, V.A., Kiselev, K.A., Ryl'kov, S.A., Sokolova, V.P., 1992.Experimental processing of Vorkutinsk manganese ores by hydro-metallurgical methods. Obogashchenie Rud (Sankt-Peterburg,Russian Federation) (6), 17–18.

Arsent'ev, V.A., Yavorskaya, G.M., Kovaleva, O.V., 1991. Manganeserecovery from oxide ores by leaching with acid. SU Patent No.1624038.

Bartolozzi, M., Braccini, G., Marconi, P.F., Bonvini, S., 1994.Recovery of zinc and manganese from spent batteries. Journal ofPower Sources 48 (3), 389–392.

Binsfield, M.C., Pesic, B., Storhok, V.C., 1996. Recovery of zincand manganese from Bunker Hill Water Treatment Plant sludge: partII — electrowinning process development. In: Ramachandran, V.,Nesbitt, C.C. (Eds.), Proceedings of the 2nd International Sympo-sium on Extraction and Processing for the Treatment and Minimisa-tion of Wastes, Minerals. Metals & Materials Society, pp. 719–729.

Brantley, F.E., Rampacek, C., 1968. Manganese and iron recoveryfrom leach solutions. US Patent No. 3397130.

Calvert, J.B., 2004. Date accessed: 29/9/2005 Chromium andManganese. http://www.du.edu/~jcalvert/phys/chromang.htm.

Chen, H., Fu, C., Zheng, D., 1992. Reduction leaching of manganesenodules by nickel matte in hydrochloric acid solution. Hydromet-allurgy 28 (2), 269–275.

Choi, K.S., Han, O.H., Shin, B.S., 1996. Reduction leaching ofmanganese nodules with sodium sulfite in ammonium chloridesolution. Shigen to Sozai 112 (2), 117–121.

Coleman, T.W., Griffin, R.A., 1984. Electrodepositing metallicmanganese. US Patent No. 4478697.

Comba, P., Lei, K.P.V., Carnahan, T.G., 1991. Calcium fluoride-enhanced hydrochloric acid leaching of a manganese-bearingsilicate ore. U.S. Bureau of Mines Report of Investigations.

Cooper, H.S., Schaefer, J.C., Schmidt, E.C., 1959. Recovery ofmanganese from metallurgical by-products by chlorination. USPatent No. 2877110.

Corathers, L.A., 2004. Date accessed: 21/6/2007. U.S. Geological Survey(USGS) Minerals Yearbook: Manganese -2004, http://minerals.usgs.gov.

Corathers, L.A., 2005. Date accessed: 21/6/2007. U.S. Geological Survey(USGS) Minerals Yearbook: Manganese -2005, http://minerals.usgs.gov.

Das, R.P., 2006. Processing ofAfricanHeterogenite at Rubamin Limited,India. Paper presented at Mineral Processing 2006, Cape Town, 3–4August. South African Institute of Mining and Metallurgy.

Das, P.K., Anand, S., Das, R.P., 1998. Studies on reduction of man-ganese dioxide by (NH4)2SO3 in ammoniacal medium. Hydro-metallurgy 50 (1), 39–49.

Das, R.P., Anand, S., Das, S.C., Jena, P.K., 1986. Leaching of manga-nese nodules in ammoniacal medium using glucose as reductant.Hydrometallurgy 16, 335–344.

http://www.hitec-energy.com.au

http://www.du.edu/~jcalvert/phys/chromang.htm

http://minerals.usgs.gov





Das, S.C., Anand, S., Das, R.P., Jena, P.K., 1989. Sulfuric acidleaching of manganese nodules in the presence of charcoal. TheAusIMM Bulletin and Proceedings 294 (1), 73–76.

Das, S.C., Sahoo, P.K., Rao, P.K., 1982. Extraction of manganese fromlow-grade manganese ores by ferrous sulfate leaching. Hydromet-allurgy 8 (1), 35–47.

Das, S.C., Sahoo, P.K., Rao, P.K., Jena, P.K., 1978. Recovery ofmanganese value from ferromanganese slag through acid leaching.Transactions of the Indian Institute of Metals 31 (4), 265–267.

Das, S.C., Sahoo, P.K., Rao, P.K., 1979. Technical note on recoveryof manganese from low manganese content ferromanganese slag.Proceedings - Australasian Institute of Mining and Metallurgy,272, 25–27.

Dean, R.S., Fox, A.L., Beck, A.E., 1942. Nitrogen dioxide process forrecovery of manganese from ores. U.S. Bureau of Mines Reportof Investigations, vol. 3626. 30 pp.

Debrodt, H., Kluger, P., Lieberoth, D., 1986. Anode for electro-chemical processes. ZA Patent No. 8602688.

Demirbas, A., 1999. Non-isothermal leaching kinetics of braunite inwater saturated with sulfur dioxide. Resources Conservation andRecycling 26 (1), 35–42.

Drinkard, Jr., W.F., Woerner, H.J., 1997. Leaching of metallurgicaldust wastes in nitric acid for recovery of metal values. WO PatentNo. 9716230.

Dry, M.J., Irio, G., Jacobs, D.F., Cole, P.M., Feather, A.M., Sole, K.C.,Engelbrecht, J., Matchett, K.C., Cilliers, P.J., O'Kane, P.T.,Dreisinger, D.B., 1998. Cu/Co tailings treatment project, Demo-cratic Republic of Congo. Proceedings of ALTA Nickel/ CobaltPressure Leaching and Hydrometallurgy Forum. ALTA Metallur-gical Services, Melbourne, Australia. 36 pp.

Dundua, R., Agniashvili, G., 1999. Manganese recovery from residualslimes in the electrochemical manufacture of manganese dioxide.Izvestiya Akademii Nauk Gruzii, Seriya Khimicheskaya 25 (1–2),151–154.

Elsherief, A.E., 2000. A study of the electroleaching of manganese ore.Hydrometallurgy 55 (3), 311–326.

Fink, C.G., Kolodney, M., 1939. Anodes for the electrowinning ofmanganese, vol. 76. Transactions of the Electrochemical Society.25 pp.

Freitas, L.R., Amaral, J.C., Mendonca, C.F., 1993. Sulfation of carajasmanganese ore with gaseous SO2. Transactions of the Institutionof Mining and Metallurgy Section C - Mineral Processing andExtractive Metallurgy 102, C130–C131.

Fuller, H.C., Edlund, V.E., Sterner, J.W., McDermaid, J.F., 1964.Metallurgical studies of rhodonite ores, Silverton district, Color-ado. III. Melting, quenching, and acid leaching of concentrates andelectrolytic recovery of manganese from solution. U.S. Bureau ofMines Report of Investigations RI 6442 (5) 31 pp.

Furlani, G., Pagnanelli, F., Toro, L., 2006. Reductive acid leachingof manganese dioxide with glucose: Identification of oxidationderivatives of glucose. Hydrometallurgy 81 (3–4), 234–240.

Gaprindashvili, V.N., Bagaturiya, L.V., Bardavelidze, M.V.,Tsveniashvili, V.S., Dzhanimanov, T.B., Gogichadze, L.D., 1993.Leaching of oxidized manganese ores in mixtures with pyrite. SUPatent No. 1792995.

Godsey, A.R. Jr., 1955. Anode for electrowinning manganese. USPatent No. 2723230.

Grimanelis, D., Neousyngouna, P., Vazarlis, H., 1992. Leaching of arich Greek manganese ore by aqueous-solutions of sulfur-dioxide.Hydrometallurgy 31 (1–2), 139–146.

Hammerquist, W.L., 1948. Electrolyte purification in manganeseelectrowinning. US Patent No. 2445826.

Hammerquist, W.L., 1951. Manganese electrowinning process. USPatent No. 2559874.

Harris, M., Meyer, D.M., Auerswald, K., 1977. The production ofelectrolytic manganese in South Africa. Journal of the SouthAfrican Institute of Mining & Metallurgy 77 (7), 137–142.

Ilea, P., Popescu, I.-C., Urda,M., Oniciu, L., 1997. The electrodepositionof manganese from aqueous solutions of MnSO4. IV: electrowinningby galvanostatic electrolysis. Hydrometallurgy 46 (1–2), 149–156.

Ismail, A.A., Ali, E.A., Ibrahim, Ibrahim, A., Ahmed, M.S., 2004. Acomparative study on acid leaching of low grade manganese oreusing some industrial wastes as reductants. Canadian Journal ofChemical Engineering 82 (6), 1296–1300.

Jacobs, J.H., Churchward, P.E., 1948. Electrowinning of manganesefrom chloride electrolytes. Journal of the Electrochemical Society94, 108–121.

Jakuszewski, B., Ignaczak, M., Dziegiec, J., Grzejdziak, A., 1972.Electrochemical oxidation and reduction on manganous sulfatein aqueous solutions. Societatis Scientiarum Lodziensis, ActaChimica 17, 83–90.

Jana, R.K., Pandey, B.D., Premchand, 1999. Ammoniacal leaching ofroast reduced deep-sea manganese nodules. Hydrometallurgy 53 (1),45–56.

Javorek, P., 1988. Ammonium carbonate leaching of nodules in man-ganese recovery. CS Patent No. 256130.

Jiang, T., Yang, Y.B., Huang, Z.C., Qiu, G.Z., 2003. Simultaneousleaching of manganese and silver from manganese-silver ores atroom temperature. Hydrometallurgy 69 (1–3), 177–186.

Kane, W.S., Cardwell, P.H., 1976. Two-stage selective leach-ing of metal values from ocean floor nodules. US Patent No.3930974.

Kaneko, T., Matsuzaki, T., Kugimiya, T., Ide, K., Kumakura, M.,Kasama, A., 1993. Improvement of Mn yield in less slag blowingat Bof by use of sintered manganese ore. Tetsu to Hagane-Journalof the Iron and Steel Institute of Japan 79 (8), 941–947.

Kanungo, S.B., Das, R.P., 1988. Extraction of metals from manganesenodules of the Indian ocean by leaching in aqueous solution ofsulfur dioxide. Hydrometallurgy 20, 135–146.

Kasimov, A.M., Slotvinskii-Sidak, N.P., Mailyan, N.A., Potapov, V.I.,1990. Roasting and leaching of manganese-containing slag. SUPatent No. 1581762.

Kawahara,M.,Mitsuo, T.,Katayama,K.I., 1991. Leaching ofmanganesenodules using sulfur dioxide as a reductant. Journal of theMining andMaterials Processing Institute of Japan 107 (12), 871–876.

Kholmogorov, A.G., Patrushev, V.V., Pashkov, G.L., Kononov, Y.S.,Mikhlina, E.V., 1998. Reductive leaching of manganese andsulfide concentrates in sulfuric acid solutions. Zhurnal PrikladnoiKhimii (Sankt-Peterburg) 71 (11), 1890–1892.

Kohga, T., Imamura, M., Takahashi, J., Tanaka, N., Nishizawa, T.,1995. Recovering iron, manganese, copper, cobalt, and high-puritynickel from sea nodules. JOM (USA) 47 (12), 40–43.

Kuh, S.E., Kim, J.W., Kim, D.S., Choi, K.S., 2001. Extraction ofmetal values frommanganese nodules in the ocean environment usingzincmatte as a reductant. Environmental Technology 22 (8), 881–887.

Kukula, M., Sikora, B., 1977. Hydrometallurgical recovery of man-ganese from slags. II. Purification of solution and electrolysis.Rudy i Metale Niezelazne 22 (8), 411–417.

Li, J., 2000. Effects of several factors on simultaneous leaching of zincsulfide ore concentrates and manganese ores. Guizhou GongyeDaxue Xuebao, Ziran Kexueban 29 (3), 10–14.

Li, X., Zhou, Q., Yang, J., Jiang, X., 1997. Reduction leaching ofocean poly metallic nodules with dilute hydrochloric acid solution.Kuangye Gongcheng 17 (2), 50–52.


Lindermann, W., Dombrowsky, C., Sewing, D., Muller, M., Engel, S.,Joppien, R., 1994. The BATENUS process for recycling batterywaste. In: Harris, B., Krause, E. (Eds.), Proceedings of ImpurityControl and Disposal in Hydrometallurgical Processes- 24thAnnual Hydrometallurgical Meeting, CIM, Canadian Inst. Mining,Metallurgy and Petroleum, pp. 197–204.

Louis, P., Martin, J.P., 1976. Increased current efficiency duringthe electrodeposition of manganese from sulfate electrolytes,Report - National Institute for Metallurgy (South Africa),vol. 1859. 27 pp.

Lu, H., Zou, X., 2001. A comprehensive utilization process for blackmanganese-silver ores by pyrite reducing method. Rare Metals(Beijing, China) 20 (3), 142–146.

Mantell, C.L., 1944. Anodes for the electrowinning of Mn. US PatentNo. 2340400.

Mantell, C.L., Ferment, G.R., 1969. Electrowinning manganese. USPatent No. 3455799.

Mantell, C.L., Hammerquist, W.L., 1944. Mg control in Mn elec-trowinning. US Patent No. 2348742.

Maslenitskii, N.N., Mil'ner, R.S., Belikov, V.V., 1969. Laboratorystudy of dithionate treatment of three samples of low-grade man-ganese slimes. Obogashchenie Rud (Sankt-Peterburg, RussianFederation) 14 (2), 45.

Mittal, N.K., Sen, P.K., 2003. India's first medium scale demonstrationplant for treating poly-metallic nodules.Minerals Engineering 16 (9),865–868.

Mohanty, P.S., Ghosh, M.K., Anand, S., Das, R.P., 1994. Leachingof manganese nodules in ammoniacal medium with elementalsulfur as reductant. Transactions of the Institution of Miningand Metallurgy Section C - Mineral Processing and ExtractiveMetallurgy, 103, C151–C155.

Mukherjee, A., Raichur, A.M., Natarajan, K.A., 2004. Recentdevelopments in processing ocean manganese nodules-a criticalreview. Mineral Processing & Extractive Metallurgy Review 25,91–127.

Naik, P.K., Das, S.C., Sukla, L.B., 2002. Extraction of manganesefrom low-grade Nishikhal ore using pyritiferous lignite in acidicmedium. Minerals & Metallurgical Processing 19 (2), 110–112.

Naik, P.K., Nathsarma, K.C., Das, S.C., Misra, V.N., 2003. Leaching oflow-grade Joda manganese ore with sulfur dioxide in aqueousmedium. Transactions of the Institutions of Mining and Metallurgy,Section C: Mineral Processing and Extractive Metallurgy 112 (2),C131–C134.

Naik, P.K., Sukla, L.B., Das, S.C., 2000. Aqueous SO2 leachingstudies on Nishikhal manganese ore through factorial experiment.Hydrometallurgy 54 (2–3), 217–228.

O'Keefe, T.J., 1984. Techniques for evaluating electrolytes for metalrecovery. Journal of Electroanalytical Chemistry and InterfacialElectrochemistry 168 (1–2), 131–146.

Omarov, N.S., Beisembaev, B.B., 1996. Use of pyrite flotation concen-trate as a reducing agent for manganese leaching from oxidizedores. Kompleksnoe Ispol'Zovanie Mineral'Nogo Syr'Ya (6),82–85.

Osawa, K., 1973. Electrolytic winning of manganese metal. JP PatentNo. 48042767.

Pagnanelli, F., Sambenedetto, C., Furlani, G., Vegliò, F., Toro, L.,2007. Preparation and characterisation of chemical manganesedioxide: effect of the operating conditions. Journal of PowerSources 166, 567–577.

Pahlman, J.E., Khalafalla, S.E., 1979. Selective recovery of nickel,cobalt, manganese from sea nodules with sulfurous acid. US PatentNo. 4138465.

Pahlman, J.E., Khalafalla, S.E., 1988. Leaching of domestic man-ganese ores with dissolved sulfur dioxide. U.S. Bureau of MinesReport of Investigations, Twin Cities, US.

Pahlman, J.E., Rhoades, C.A., Chamberlain, P.G., 1987. Dual leachingmethod for recovering silver and manganese from domesticmanganiferous silver deposits. U.S. Bureau of Mines Report ofInvestigations, Twin Cities, US.

Paixao, J.M.M., Amaral, J.C., Memoria, L.E., Freitas, L.R., 1995.Sulfation of Carajas manganese ore. Hydrometallurgy 39 (1–3),215–222.

Parissis, G., Winand, R., 1992. Manganese electrodeposition fromchloride, sulfate and mixed solutions, including high currentdensities, ATB Metallurgie 32 (1–2), 16–19, 22–8.

Partenov, D., Stefanova, V., Avramov, A., Chimbulev, M., 2004.Kinetics of leaching of polymetallic concretions in an aqueoussolution of SO2. Izvestiya Vysshikh Uchebnykh Zavedenii,Tsvetnaya Metallurgiya (1), 15–19.

Pesic, B., Storhok, V.C., Binsfield, M.C., 1996. Process for recoveryof Zn and Mn from Bunker Hill Water Treatment Plant sludge:Part I — leaching and solution purification. In: Ramachandran,V., Nesbitt, C.C. (Eds.), Proceedings of the 2nd InternationalSymposium on Extraction and Processing for the Treatment andMinimisation of Wastes, Minerals, Metals & Materials Society,pp. 699–710.

Petrie, L.M., 1995. Molecular interpretation for SO2 dissolutionkinetics of pyrolusite. manganite and hematite, Applied Geochem-istry 10 (3), 253–267.

Pisarczyk, K., 1995. Manganese and manganese alloys, 4th Edition.Kirk-Othmer Encyclopedia of Chemical Technology, vol. 15. JohnWiley & Sons, New York.

Popov, S.A., Klushin, V.N., Rodionov, A.I., Makotkin, A.V., Maskova,T.O., 1981. Study of equilibrium characteristics of a sulfur dioxide-air-aqueous suspension of manganese dioxide-based industrial by-product system. Trudy Instituta - Moskovskii Khimiko-Tekhnolo-gicheskii Institut Imeni D. I. Mendeleeva 119, 24–27.

Ptitsyn, A.N., Galkova, L.I., Ledvii, V.V., Dobyshev, B.V., Skopov, S.V.,2001a. Electrochemical manufacture of manganese dioxide fromleaching of lean Mn-carbonate ores. RU Patent No. 2172791.

Ptitsyn, A.N., Galkova, L.I., Ledvii, V.V., Skopov, S.V., 2001b.Recovery of metal values from chemical byproducts of spent Mn-Zn electric batteries. RU Patent No. 2164955.

Raghavan, R., Upadhyay, R.N., 1999. Innovative hydrometallurgicalprocessing technique for industrial zinc and manganese processresidues. Hydrometallurgy 51 (2), 207–226.

Rappas, A.S., 1985. Processes relating to the extraction of metals fromlaterites and ocean manganese nodules, Official Gazette, 8 Oct.

Ravitz, S.F., Wyman, W.F., Back, A.E., Tame, K.E., 1946. Thedithionate process for recovery of manganese from low-grade ores.American Institute of Mining Metallurgical Engineers MetalsTechnology 13 (No. 6, Tech. Pub. No. 2064) 10 pp.

Rodriguez, M., Rivarola, J., Ruiz, M.D., 2004. The effects ofcarboxylic acid addition on hydrofluoric acid autoclave leachingof a ferrocolumbite. Hydrometallurgy 74 (1–2), 39–46.

Rokukawa, N., 1990. Recovery of nickel, cobalt, and copper fromocean manganese nodules with mixed solution of ammoniumcarbonate and ammonium sulfite. Shigen to Sozai 106 (4),205–209.

Rolf, R.F., 1969. Selective recovery of manganese and iron from ores.US Patent No. 3471285.

Saha, A.K., Khan, Z.H., Akerkar, D.D., 1993. Reduction of man-ganese sea nodules with noncoking coal. Transactions of the IndianInstitute of Metals 46 (2), 63–69.


Sahoo, R.N., Naik, P.K., Das, S.C., 2001. Leaching of manganese fromlow-grade manganese ore using oxalic acid as reductant in sulfuricacid solution. Hydrometallurgy 62 (3), 157–163.

Sanigok, U., Bayramoglu, M., 1988. Bench-scale manganese sulfateproduction from low-grade pyrolusite ores. - Part III. Chimica ActaTurcica 16 (1), 9–20.

Shibata, J., Baba, Y., 1995. Development of recovery process ofvaluable metals from waste lithium battery. Minerals, Metals andMaterials Society/AIME (USA) 257–261.

Sole, K.C., Feather, A.M., Cole, P.M., 2005. Solvent extraction insouthern Africa: an update of some recent hydrometallurgicaldevelopments. Hydrometallurgy 78 (1–2), 52–78.

Suliakas, A., Yanitskii, I.V., Juseviciene, G., 1973. Behavior of seleniteor selenate additions of an anode during electrodeposition ofmanganese. Lietuvos TSR Mokslu Akademijos Darbai, Serija B:Chemija, Technika, Fizine Geografija (4), 47–54.

Sventsitskii, A.T., Nosenkov, A.N., Trunev, S.V., Dmitrevskii, B.A.,Treushchenko, N.N., Yur'eva, V.I., Ivanova, N.Y., 2003. Acidicredox leaching of lean manganese ores, slimes, and dust fromferroalloy furnaces. RU Patent No. 2213155.

Szabo, L.J., 1976. Recovery of metal values from manganese deep seanodules using ammoniacal cuprous leach solutions. US Patent No.3983017.

Taylor, T., 1993. Electrolytic manganese dioxide production byAustralian Manganese Company Limited, Newcastle. AustralasianMining and Metallurgy, Clunies Ross House, Victoria, Australia2 (Monograph No.19), 1237–1238.

Thomas, G., Whalley, B.J.P., 1958. The leaching of manganese frompyrolusite ore by pyrite. Canadian Journal of Chemical Engineer-ing 36, 37–43.

Veglio, F., Beolchini, F., Gasbarro, A., Toro, L., Ubaldini, S.,Abbruzzese, C., 1997. Batch and semi-continuous tests in thebioleaching of manganiferous minerals by heterotrophic mixedmicroorganisms. International Journal of Mineral Processing 50 (4),255–273.

Veglio, F., Toro, L., 1994a. Reductive leaching of a concentratemanganese-dioxide ore in acid-solution — stoichiometry and pre-liminary kinetic-analysis. International Journal of Mineral Proces-sing 40 (3–4), 257–272.

Veglio, F., Toro, L., 1994b. Fractional factorial-experiments in thedevelopment of manganese-dioxide leaching by sucrose insulfuric-acid-solutions. Hydrometallurgy 36 (2), 215–230.

Vu, H., Jandova, J., Lisa, K., Vranka, F., 2005. Leaching of manganesedeep ocean nodules in FeSO4–H2SO4–H2O solutions. Hydromet-allurgy 77 (2005), 147–153.

Wanamaker, E.M., Hammerquist, W.L., Morgan, Jr., W.D. 1951.Molybdenum control in manganese electrowinning. US Patent No.2538995.

Wanamaker, E.M., Morgan, Jr., W.D. 1943. Catholyte purification inthe electrowinning of Mn. US Patent No. 2325723.

Ward, C., Cheng, C.Y., Urbani, M.D., 2004. Manganese - from wasteto high-tech material. Publications of the Australasian Institute ofMining and Metallurgy, pp. 241–246. 2/2004.

Ward, C.B., 2005. Acidic leaching of manganese from lean oxide oreswith extraction stage for purity. WO Patent No. 2005012582.

Welsh, J.Y., Sochol, I., 1978. Recovering metal values from marinemanganese nodules, Official Gazette, 31 Oct.

Welsh, J.Y., 1973. Recovering high-purity, free-flowing, crystallinemanganese dioxide from impure manganese nitrate solutions. USPatent No. 3780158.

Welsh, J.Y., 1978. Recovering manganese dioxide crystals from man-ganese nitrate solutions. RO Patent No. 63988.

Yaozhong, L., 2004. Laboratory study: simultaneous leaching silver-bearing low-grade manganese ore and sphalerite concentrate.Minerals Engineering 17 (9–10), 1053–1056.

Yavorskaya, G.M., Arsentev, V.A., Kucher, V.L., Kovaleva, O.V.,1992. Redox leaching of manganese from oxide feed materials.SU Patent No. 1733492.

Yoon, B.H., Kim, D.R., 1986. Recovery of manganese from electro-lytic zinc anodic slime and from scrap dry cells. Kumsok PyomyonCholi 19 (1), 13–19.

Yucel, O., Ari, M.E., 2001. Carbothermic smelting of Tavas manganeseore. Turkey, High Temperature Materials and Processes 20 (5–6),345–351.

Zhang, W., Cheng, C.Y., 2006. Manganese metallurgy, recovery andcontrol — a literature review. Part I: Manganese MetallurgyCSIRO Minerals, Australia, DMR - 2809.

Ziyadanogullari, R., Buyuksahin, M., 1995. Recovery of MnSO4 fromlow-grade pyrolusite ores, and of MnSO4 and silver from man-ganese-silver ores. Separation Science and Technology 30 (3),477–486.

Manganese Metallurgy Part I

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