Hydrometallurgy 98 (2009) 1–9

Contents lists available at ScienceDirect

Hydrometallurgy

j ourna l homepage: www.e lsev ie r.com/ locate /hydromet

A literature review of the recovery of molybdenum and vanadium from spenthydrodesulphurisation catalystsPart I: Metallurgical processes

Li Zeng, Chu Yong Cheng ⁎Parker Centre for Integrated Hydrometallurgy Solutions/CSIRO Minerals, PO Box 7229, Karawara, WA 6152, Australia

⁎ Corresponding author. Tel.: +61 8 9334 8916; fax: +E-mail address: chu.cheng@csiro.au (C.Y. Cheng).

0304-386X/$ – see front matter. Crown Copyright © 20doi:10.1016/j.hydromet.2009.03.010

a b s t r a c t

a r t i c l e i n f o

Article history:Received 20 January 2009Received in revised form 23 March 2009Accepted 23 March 2009Available online 31 March 2009

Keywords:MolybdenumVanadiumSpent hydrodesulphurisation catalystLeaching

Catalysts are widely used in petroleum refining and chemical industries. Among secondary resources, spentcatalysts are undoubtedly very important because of not only their large amounts and enormous economicvalues, but also because of the environmental concerns if disposed off. Spent hydrodesulphurisation catalystsusually consist of molybdenum sulphide mixed with sulphides of vanadium, cobalt and nickel on an aluminacarrier. A wide variety of metallurgical processes are used to treat these catalysts. The processes vary in theirselectivity for metals and complexity of operation, but adopt one of the following approaches: 1. acidleaching with either H2SO4, HCl or (COOH)2, often after roasting; 2. caustic leaching with NaOH, sometimesafter roasting; 3. salt roasting with Na2CO3, NaCl or NaOH followed by leaching with water or Na2CO3;4. smelting either directly or after calcination; 5. anhydrous chlorination; 6. bioleaching.Roasting followed by sulphuric acid leaching seems to be the best option since all of the valuable metalsdissolve. However the downstream processes to produce separate products with high purity are relativelycomplex. Sodium carbonate roasting followed by water leaching is a good option since molybdenum andvanadium are selectively extracted over aluminium, nickel and cobalt. Bioleaching offers good prospects forrecovering valuable metals and at the same time, generates much less environmental pollution. However,much more research work is needed before it can be commercialised.After leaching, the metals in leach solutions have to be separated and purified by conventional separationtechniques such as precipitation, adsorption, ion exchange and solvent extraction. Part II of this reviewconsiders the application of these methods, especially, solvent extraction for treating such leach solutions.

Crown Copyright © 2009 Published by Elsevier B.V. All rights reserved.

Contents

1. Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 22. Metallurgical processes for treating spent HDS catalysts . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2

2.1. Acid leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 22.1.1. Roasting–leaching with concentrated sulphuric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 32.1.2. Roasting–leaching with dilute sulphuric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 32.1.3. Roasting–pressure leaching with sulphuric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 32.1.4. Direct leaching with sulphuric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 32.1.5. Direct leaching with hot hydrochloric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42.1.6. Retorting–leaching with concentrated hydrochloric acid . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42.1.7. Leaching with oxalic and citric acids . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 4

2.2. Caustic leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42.2.1. Roasting–caustic leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42.2.2. Direct caustic leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 52.2.3. NaOH/NaAlO2 leaching under pressure. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 52.2.4. Caustic–sulphuric acid leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5

2.3. Smelting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 52.3.1. Direct smelting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5

61 8 9334 8001.

09 Published by Elsevier B.V. All rights reserved.

2 L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

2.3.2. Calcining and smelting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 52.4. Selective anhydrous chlorination . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 5

2.4.1. Roasting–chlorination . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 62.4.2. Direct chlorination . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 6

2.5. Salt roasting–leaching and salt leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 62.5.1. Oxidative roasting–NaCl/water vapour roasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 62.5.2. Sodium carbonate roasting and water leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 72.5.3. Direct sodium carbonate leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 7

2.6. Bioleaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 73. Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8Acknowledgments . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8References. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 8

Table 1Compositions of different kinds of catalysts (%) (based on Zhang and Zhao, 2005).

Catalyst Mo Fe2O3 NiO Co V2O5 Al2O3 Cu Si P C+S Others

Mo–Fe 5 48 – – – 15 1 – – 30 1Mo–Ni 18.5 – 4.0 – – 52.5 – 4.7 2 – 18Mo–Co 8.14 – 3.2(Ni) 1.94 – 24.3 – – – 25.37 37.02Mo–V 1–10 – 1–12 – 1–15 – – – – 2–12 1–40

– Unknown or negligible.

1. Introduction

Molybdenum and vanadium are important metals in human lifeand industry. In recent years, the world molybdenum and vanadiumdemands have been driven by soaring steel production, particularly inChina. As typical transition elements, molybdenum and vanadiumhave wide applications in the steel industry as alloying elements.Their extensive consumption in all industrialised countries isattributed to the numerous beneficial effects they impart on ironand steel (Kar et al., 2004). In addition, molybdenum is also widelyused as catalysts in the petrochemical industry and vanadium inchemical and aerospace industries because of their many superiorperformances (Anon, 2003). The largest usage of molybdenumcatalysts is in the desulphurisation of petroleum, petrochemicalsand coal-derived liquids to minimise sulphur dioxide emission fromfuel combustion. Spent catalysts from hydrogenation plants are themajor secondary sources of molybdenum. Although many countriescurrently recycle a large amount of spent hydrodesulphurisationcatalysts, more efficient and effective recycling methods are requiredto be developed.

With the rapid growing demand for molybdenum and vanadiumproducts and ceaseless exploitation of all the mineral resources ofthese twometals in theworld, primarymineral supplies are becomingmore and more insufficient to meet demands, resulting in increasingexploitation of secondary sources. Among secondary resources, spentcatalysts are undoubtedly very important. Hydrodesulphurisation(HDS) catalysts account for about one third of the total worldwidecatalyst consumption and are used for sulphur removal in petroleumrefineries. Spent HDS catalysts contain molybdenum, cobalt, nickeland/or vanadium on an alumina carrier. These catalysts are regardedas the most important catalysts for recycling these metals due to notonly their large amounts andmetal values, but also the environmentalconcerns if disposed off (Furimsky, 1996). In recent years, variouspyro-metallurgical and hydrometallurgical processes have beenstudied and developed for recovery of these metal values fromspent HDS catalysts. Among these processes, hydrometallurgicalprocesses are the most important ones. Therefore, in this part of thereview, recent developments in the leaching of spent HDS catalysts arereviewed in detail.

2. Metallurgical processes for treating spent HDS catalysts

HDS catalysts account for about one third of the total worldwidecatalyst consumption and are mostly used for the desulphurisation ofvarious oil fractions and also for the removal of metal impurities. Thefresh catalysts mostly consist of molybdenum oxide mixed mainlywith oxides of vanadium, cobalt or nickel on an alumina carrier. Thecompositions of some kinds of catalysts are shown in Table 1. Duringoperation, trace amounts of vanadium and nickel impurities in thecrude oil gradually deposit onto the catalysts. The operating condi-tions are favourable for the formation of metal sulphides, such as

sulphides of molybdenum, cobalt, nickel and vanadium (Biswas et al.,1985; Jong and Rhoads, 1989). Catalysts can also be contaminated bycarbonaceous deposits. Spent HDS catalysts generally consist of 10–30%molybdenum,1–12% vanadium, 0.5–6% nickel, 1–6% cobalt, 8–12%sulphur, 10–12% carbon and the balance is alumina, which makes iteconomically viable for recovery of valuable metals (Biswas et al.,1986). The composition of spent catalysts varies depending on howmuch vanadium and nickel are contained in the crude oil. Heavy oilfractions contain more vanadium and nickel while others containnegligible vanadium.

During the past decade, much research work was conducted torecover molybdenum, vanadium and other valuable metals fromspent HDS catalysts (Chen et al., 2006). A number of processes andtechnologies for the recycling of valuable metals from spent catalystshave been developed and reported (Zhang et al., 1995, 1996; Sun et al.,1998, 2001; Van den Berg et al., 2002; Kar et al., 2005). Basically, thereare two main processes, hydro- and pyro-metallurgical processes torecover metal values from spent catalysts. These processes includeroasting, acid and caustic leaching, smelting, anhydrous chlorination,bioleaching and salt roasting followed by water leaching. The mostcommon approach to convert sulphides to oxides is roasting. This alsoremoves any carbon residues or oil on the spent catalyst and iscommonly followed by other processes such as acid and causticleaching or smelting.

2.1. Acid leaching

There are two main routes to process spent catalysts in the acidleaching stage. One is direct acid leaching under pressure or hightemperature; the other is roasting followed by acid leaching with orwithout pressure and high temperature. Roasting is commonly carriedout in a temperature range of 500–700 °C in air to convert metalsulphides to oxides sincemolybdenum, vanadium, nickel and cobalt assulphides in the catalyst are difficult to dissolve. According to Ward(1989), aluminates and some other insoluble complexes formed above700 °C. In some cases, steam was used, resulting in better aluminiumextraction.

Acid leaching is commonly used if the recovery of all metals fromthe spent catalyst is required. All of the metals and usually somealumina carrier dissolve whereas there is little dissolution of silica.Sulphuric, hydrochloric and nitric acids and some organic acids suchas oxalic and citric acids are used for leaching spent catalysts.

Fig. 1. Metallurgical processes for spent HDS catalysts involving acid leaching (based on Ho, 1992).

3L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

Sulphuric and hydrochloric acid are commonly used (Rastas et al.,1983; Lee et al., 1992). Capital costs for the installation of theequipment prohibit the industrialisation of the nitric acid leachingprocess involving the use of liquid oxygen. Some metallurgicalprocesses for spent HDS catalysts involving acid leaching are shownin Fig. 1.

2.1.1. Roasting–leaching with concentrated sulphuric acidIn this leaching process, the concentrations of sulphuric acid

ranged from 30% to 70% (v/v) and the leaching temperatures werein the range of 100–200 °C (Suzuki and Gao, 1982). There arevarious ways of separating and recovering the metal values fromthe leach solution. Recovery of up to 90% metals was achieved afterprecipitation.

A process of roasting spent HDS catalysts containing molybdenum,vanadium, nickel and cobalt at 800–850 °C in a fluidised bed followedby sulphuric acid leaching was reported by Jocker (1993). This processis called the Metrex process. After roasting and leaching, solventextraction was then used to recover firstly, the molybdenum andsubsequently, the vanadium, cobalt and nickel. In the “incineration”step: a part of the molybdenum was recovered as MoO3 byevaporation and sublimation. After thermal treatment, the aluminacarrier was left insoluble, and therefore, a major aluminiumprecipitation stage was not required. The disadvantage of this process

Fig. 2. A conceptual flow sheet of the Metrex process (based on Anon., 1992).

is that it is complex and difficult to control. In addition, there is nooption for vanadium recovery. The flow sheet of the Metrex process isshown in Fig. 2.

2.1.2. Roasting–leaching with dilute sulphuric acidIn this leaching process, after roasting, dilute sulphuric acid and an

oxidant such as hydrogen peroxide were used to selectively removenickel and cobalt from molybdenum and vanadium (Mihashi et al.,1982). The spent catalyst was heated first in nitrogen at 450–500 °C,then heated to 950–1000 °C followed by leaching with a 5 g/L H2SO4

and 20 g/L H2O2 solution. All of the nickel and cobalt can be leachedinto the solution. However, 10% of the molybdenum and 35% of thevanadium also entered the solution which needed to be recovered.The residue containing the molybdenum and vanadium was stirredwith a caustic solution to leach molybdenum and vanadium to asolution containing no nickel and cobalt.

2.1.3. Roasting–pressure leaching with sulphuric acidIn this leaching process, after roasting, the spent catalyst was

leached in an autoclave at 180–220 °C using a solution containingalkali sulphates such as potassium, ammonium and/or aluminiumsulphates and sulphuric acid (Ho, 1992). The aluminium in thesolution was precipitated out in the form of alunite which could befiltered off in the residue. Solvent extraction was then used to recovermolybdenum, vanadium, nickel and cobalt.

2.1.4. Direct leaching with sulphuric acidDirect leaching spent catalysts can be carried out with concen-

trated or dilute sulphuric acids. Leaching with concentrated sulphuricacid under 1–35 atm H2S pressure was reported in a US patent (Hyatt,1987). It caused aluminium to remain in the solution but other metalsformed solid sulphides. After separation from aluminium, they werethen oxidised to sulphates under 1–35 atm O2 pressure. A Polishpatent reported a process involving fusing the spent catalyst withconcentrated sulphuric acid in 100–200 °C followed by water leaching(Grzechowiak et al., 1987). Dilute sulphuric acid leaching usually gavelow recoveries of metal values from the molybdenum–cobalt catalystsince the redox potential of the solution was not high enough tooxidise the metal sulphides. However, a 10% w/w sulphuric acidleaching followed by solvent extraction was reported to recover over90% of the molybdenum and nickel from a molybdenum–nickelcatalyst (Siemens et al., 1986).

Fig. 3. Metallurgical processes for spent HDS catalysts involving caustic leaching (based on Ho, 1992).

4 L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

2.1.5. Direct leaching with hot hydrochloric acidHydrochloric acid leaching was chosen because metals form

different chloro-complexes which created opportunities for theseparation of these metals. Moreover, the effluent solution containssodium chloride which is an environmentally acceptable substance.

Hot concentrated hydrochloric acid (N90 °C) was used to leachspent catalysts (Ward, 1989). More than 99% of the metals, includingaluminium, dissolved leaving a silica residue. Chlorine was liberateddue to vanadium(V) reduction to vanadium(IV) and the off-gas wasscrubbed with water and caustic soda to form marketable hypochlor-ite. Aluminium was crystallised out in the form of AlCl3·6H2O byinjecting hydrochloric acid gas into the solution to a concentration of450 g/L. The other metals were separated and recovered byprecipitation or solvent extraction.

2.1.6. Retorting–leaching with concentrated hydrochloric acidRetorting the spent catalysts under steam to decrease the amount

of oil followed by leaching with concentrated hydrochloric acid wasused to selectively leach vanadium, cobalt, nickel and aluminium aschlorides, leavingmolybdenum in the residue (Sefton et al., 1989). Themolybdenum was then recovered by roasting the residue andconsequently dissolving in hydrochloric acid.

A comparison between leaching with H2SO4 and HCl acids wascarried out by Biswas et al. (1986), indicating that hydrochloric acidleaching seemed to achieve higher recoveries of metals but the leachsolution was more corrosive than that in sulphuric acid leaching. Itwas found that there was little difference between hydrochloric andsulphuric acid leaching of molybdenum and cobalt provided the endpH was similar. Sulphuric acid is most commonly used because ofmore flexible material allowances for reactors, lower costs and betterrecirculation possibilities.

2.1.7. Leaching with oxalic and citric acidsOxalic acid used as a complexing agent to selectively extract

vanadium from the spent catalyst was reported by (Lee et al., 1992).The spent catalyst was firstly treated with warm toluene to removeresidual oil. After subsequent treatment with hexane and vacuumdrying, 8% oxalic acid solution was used to extract vanadium from thecatalyst. Citric acid leaching was also studied and the resultant leachsolution contained the same amount of vanadium as that resultingfrom oxalic acid leaching. However, it also contained small amounts ofmolybdenum, nickel and aluminium because citric acid was not soselective as oxalic acid for vanadium over molybdenum and nickel.This method can only be used as a part of a comprehensive process torecover all metal values since the leach residue still contained largeamount of valuable metals.

2.2. Caustic leaching

Alkali leaching can be used to selectively dissolve molybdenumand vanadium from the spent HDS catalyst. It also dissolves somealuminium but leaves nickel, cobalt and iron in the residue to somedegree. In most cases, the metal sulphides were oxidised first byroasting or leaching under pressure (Millsap and Reisler, 1978; Biswaset al., 1985; Ikeyama, 1987). Soluble sodium molybdate, vanadate andaluminatewere then formed and dissolved in the leach solution. Therewere three main leaching routes, including roasting followed bycaustic leaching, direct hot caustic leaching usually under pressureand caustic/sodium aluminate leaching (Yasuhara, 1982; Jong andSiemens, 1985; Grzechowiak et al., 1987; Wiewiorowsk et al., 1987;Crnojevich et al., 1990). Cobalt and/or nickel in the leached catalystresidue could be dissolved by acid or ammonium carbonate.Alternatively the residue could be sent to a smelter. Metallurgicalprocesses for spent HDS catalysts involving caustic leaching are shownin Fig. 3.

2.2.1. Roasting–caustic leachingThis process was used to treat spent HDS catalysts containing

vanadium as well as molybdenum, nickel and cobalt. It was reportedthat one of such catalysts containing up to 15% vanadium was firstlywashed with ligroin, then ground and roasted at 630 °C in air toremove carbon and sulphur (Biswas et al., 1986). In the process,gamma alumina was converted to alpha alumina which is not readilydissolved in the caustic solution. Leaching was carried out at theboiling point but not under pressure. With the increase in causticconcentration from 0.25 M to 2 M, the dissolution of molybdenum,vanadium and alumina increased, but the increase was considerablyfavourable for vanadium (Parkinson and Ishio, 1987). The disadvan-tages are the difficulty of separating the solid–liquid phases and thecomplexity of recovering valuable metals from the leach solution.

Villarreal et al. (1999) reported that vanadium and molybdenumwere leached with ammonia or NaOH solution as ammonium orsodium vanadate and molybdate from the spent petroleum catalyst.The spent catalyst was pre-treated by washing the CS2 solutionfollowed by heating at different temperatures in an electric oven toeliminate the carbon and sulphur residues as CO2 and SO2,respectively. CS2 is a good solvent for the elimination of organiccompounds and the recovery of sulphur from the spent catalyst toavoid the contamination of atmosphere by SO2. Such treatment didnot cause structural change of the catalyst. Therefore, this processmight be applied to the regeneration of the catalyst for re-use.Leaching with NaOH solution of 10% concentration led to theextraction of 88.5% vanadium and 91.8% molybdenum.

Fig. 5. Smelting routes for spent catalysts (based on Ho, 1992).

5L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

2.2.2. Direct caustic leachingThere are three main routes in direct leaching, dilute caustic

leaching at atmospheric pressure (Jong and Siemens, 1985), dilutecaustic leaching under pressure (Yasuhara, 1982) and concentratedcaustic leaching under pressure (Grzechowiak et al., 1987). Directleaching without pressure, especially at low caustic concentration isnot practicable because of the lower recoveries of molybdenum andvanadium (Biswas et al., 1986; Wiewiorowsk et al., 1987).

For a catalyst containing negligible vanadium and cobalt, dilutesodium hydroxide leaching carried out at atmospheric pressure and100 °C dissolved over 90% of molybdenum (Jong and Siemens, 1985).Aluminium leaching dropped dramatically from 34% to 0.06% with theincrease in solid/liquid ratio while molybdenum leaching decreasedslightly. Leaching with a 4% caustic solution in an autoclave resulted inthe dissolution of over 95% of vanadium and molybdenum for avanadium-containing catalyst with only 6% aluminium dissolution(Yasuhara, 1982).

A direct leaching of spent catalysts with concentrated causticsolution at a relatively high temperature of 177 °C and pressure wasreported (Grzechowiak et al., 1987). This method involved a largeamount of reagents and the subsequent process for recovering thevaluable metals was very complicated. Both direct hot caustic leachingunder pressure and combined roasting–caustic leaching have beenused commercially (Parkinson and Ishio, 1987).

2.2.3. NaOH/NaAlO2 leaching under pressureIn this process, the catalyst containing about 4.1% molybdenum, 7%

vanadium, 1.3% nickel, 1.4% cobalt, 7.4% sulphur and 6.3% carbon byweight was slurried with water and sodium hydroxide and/or sodiumaluminate in an amount at least stoichiometrically sufficient toconvert molybdenum to sodium molybdate, vanadium to sodiumvanadate and sulphur to sodium sulphate in the presence of oxygen(Wiewiorowsk et al., 1987). The slurry was then heated in a pressurevessel with temperatures in the range of 150–250 °C. The addition ofsodium aluminate to the leach solution reduced aluminium dissolu-tion. The sodium aluminate can be recycled. The flow sheet of the so-called CRI-MET process is shown in Fig. 4.

2.2.4. Caustic–sulphuric acid leachingPark and Mohapatra (2007) reported a caustic–sulphuric acid

leaching process. In the process, the spent catalyst was first treated byalkali solutions to remove molybdenum followed by sulphuric acidleaching, resulting in a solution rich in Ni, Co and Al. Before the two-stage alkali–acid leaching process, the spent catalyst was roasted at500 °C to remove C and S. Recoveries of up to 98% Mo, 93% Co, 90% Niand 21% Al were achieved. After roasting, caustic leaching followed bysulphuric acid leaching seemed to give high overall dissolutions ofmolybdenum, vanadium, nickel and cobalt in spent catalysts.

Fig. 4. A conceptual flow sheet of CRI-MET process (based onWiewiorowsk et al., 1987).

2.3. Smelting

Smelting is usually used to directly produce alloy from the spentcatalyst. There are two main approaches to smelting: direct smeltingand calcining–smelting (Fig. 5).

2.3.1. Direct smeltingDirect reductive smelting at 1500 to 1700 °C was reported in a few

patents which only dealt with catalysts without vanadium (Ogui et al.,1971; Krismer et al., 1979; Oshiumi, 1979; Mueller et al., 1980). Thespent catalyst not subject to oxidative roasting was reduced in thepresence of carbon and CaO to make the metals in the form of alloyswhich sunk to the bottom of the furnace and the alumina floatedabove the alloy to form an immiscible slag. Scrap iron was usuallyadded to promote alloy formation by giving more volume to themolten phase, collecting the metals, lowering the melting point of thealloy and improving slag separation (Howard and Barnes, 1991). Limewas also added to promote slag formation by lowering its meltingpoint. Recoveries of molybdenum and cobalt from this processreached 93% and 91%, respectively (Oshiumi, 1979).

2.3.2. Calcining and smeltingTwo sequential steps, calcining and smelting, were used to recover

valuable metals from spent catalysts (Howard and Barnes, 1991). Thespent catalyst was calcined first under oxidising conditions in atemperature range from 760–870 °C to remove sulphur, free carbon,water and hydrocarbons. In this step, various amounts of molybde-numwere volatilised asMoO3which could be captured in bag filters asa saleable product after cooling. In the next step, the calcine wassmelted preferably along with scrap iron at 1650–2400 °C in anelectric arc furnace. The remaining molybdenum was volatilised andcollected in the same way. A reductant, such as natural gas wasinjected to reduce the oxides of vanadium, cobalt and nickel to themetallic state, which formed an alloy. The molten alumina wasseparated from the alloy of heavy metals by gravity and removedperiodically. Fresh calcine was added until the volume of the alloyphase was large enough for tapping. The process was expected torecover about 99% of the heavy metals.

Medvedev and Malochkina (2007) presented a method ofsublimation, which was performed in a muffle electrical resistancefurnace at 1200 °C to recover Mo directly as MoO3 from spentcatalysts. With the optimisation of operating parameters, the extrac-tion of molybdenum into sublimate of no less than 80% can beobtained in one stage of sublimation.

2.4. Selective anhydrous chlorination

Selective anhydrous chlorination for the recovery of valuablemetals from spent HDS catalysts is a method where molybdenum andvanadium are volatilised as their chlorides or oxychlorides by chlorineand/or hydrogen chloride gas. Molybdenum and vanadium arerecovered by selective condensation from the vapour phase whilecobalt and nickel chlorides are extracted from the chlorination residueby leaching with water or acidic solution.

Fig. 6. A conceptual flow sheet of roasting–chlorination for recycling a Ni–Mo–Alcatalyst (based on Jong and Siemens, 1985).

6 L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

2.4.1. Roasting–chlorinationSpent catalysts were usually roasted first in a temperature range of

300–500 °C to convert the metal sulphides to metal oxides and toremove moisture. The converted catalyst was then chlorinated in atemperature range of 200–500 °C. This resulted in volatilisation ofmolybdenum, vanadium and some aluminium in the form ofmolybdenum and vanadium chlorides or oxychlorides and aluminiumchloride, respectively (Gaballah and Djona, 1994). The nickel andcobalt chlorides stayed in the catalyst residue because of their highervapour pressure and then were leached out with water, leaving analumina residue. An advantage of the chlorination process is thatmolybdenum and vanadium can be separated from nickel and cobalt,which makes the subsequent processes easier. A potential processflow sheet developed for the Ni–Mo–Al catalyst using chlorine gas(Jong and Siemens, 1985) is shown in Fig. 6.

2.4.2. Direct chlorinationThe carbon, sulphur and hydrocarbons present in the raw and un-

roasted spent catalysts could be used as reducing agents and sourcesof energy. In addition, the roasting step may increase the capital andoperating costs of the process and lead to physicochemical modifica-tions of the catalyst, resulting in lower extraction of the valuableelements during chlorination. Gaballah and Djona (1995) investigateddirect selective chlorination of raw and un-roasted spent catalysts

Fig. 7. A proposed flow sheet for the extraction of valuable elements from spe

with Cl2/air, Cl2/N2 and Cl2/CO/N2 gas mixtures for the recovery ofvaluable elements. The results showed that it is possible to recovermore than 90% of Ni and Co, about 99% of Mo and up to 75% of V usinga Cl2/air gasmixture in temperatures lower than 600 °C. The proposedflow sheet is shown in Fig. 7.

2.5. Salt roasting–leaching and salt leaching

Salt roasting followed by water leaching is used to selectivelydissolve molybdenum and vanadium over aluminium, cobalt andnickel in spent catalysts. Thus, there is no significant amount ofaluminium to interfere with the separation of molybdenum andvanadium through co-precipitation (Parkinson and Ishio, 1987).Molybdenum and vanadium were converted to soluble sodium saltssuch as Na2MoO4 or Na2Mo2O7 and NaVO3 which can be leached withwater after roasting. There are two approaches for roasting: oxidisingroasting followed by salt (NaCl) roasting and direct salt roasting.

2.5.1. Oxidative roasting–NaCl/water vapour roastingIn this process, calcination of the spent catalysts in air to remove

carbon and oil and oxidation of metal sulphides to their oxides werecarried out. The oxides were then converted to soluble salts by driedsodium chloride–water vapour roasting (Biswas et al., 1985). Theprocess may include the following reactions:

NaCl þ H2OðgÞ ¼ NaOH þ HClðgÞ ð1Þ

V2O5 þ 2NaCl þ H2OðgÞ ¼ 2NaVO3 þ 2HClðgÞ ð2Þ

2MoO3 þ 2NaClþ H2OðgÞ ¼ Na2Mo2O7 þ 2HClðgÞ ð3Þ

Al2O3 þ 2NaCl þ H2OðgÞ ¼ 2NaAlO2 þ 2HClðgÞ: ð4Þ

The hydrolysis of sodium chloride is more rapid above its meltingpoint of 800 °C and, therefore, a temperature of 850 °C was used. Aninert atmosphere also favours the salt roasting reactions and nitrogenwas used as the carrier gas. A water vapour pressure of 0.245 atmwaschosen. The disadvantage of this method is that some molybdenum,vanadium, nickel and cobalt were lost probably due to the formationof volatile chlorides (Biswas et al., 1985). Leaching was carried out atthe boiling point of the solution because metal dissolution increasedwith the temperature. During leaching, aluminiumwas precipitated asaluminium hydroxide, leaving little aluminium in the solution. Therecoveries of molybdenum and vanadium reached 76% and 77%,respectively. The low metal recoveries were most probably due to the

nt catalysts by direct chlorination (based on Gaballah and Djona, 1995).

7L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

loss of metals during roasting and the incomplete dissolution ofvanadium during leaching.

2.5.2. Sodium carbonate roasting and water leachingAlthough sodium hydroxide, sodium bicarbonate and sodium

sulphate were used in direct salt roasting, sodium carbonate was themost usual salt for this process (Naka, 1972; Toda, 1989a,b).Molybdenum and vanadium sulphides were converted to correspond-ing soluble sodium salts as expressed in the following reactions:

MoS2 þ 3Na2CO3 þ 9=2O2 ¼ Na2MoO4 þ 2Na2SO4 þ 3CO2 ð5Þ

V3S4 þ 17Na2CO3 þ 39=2O2 ¼ 6Na3VO4 þ 8Na2SO4 þ 17CO2: ð6Þ

Alumina, nickel and cobalt oxides remained in the residue. Sodiumcarbonate roasting was carried out in 650–900 °C at a Na2CO3/(Mo+V) molar ratio of about 2 (Toda et al., 1979; Toda and Morimoto, 1983;Toda, 1989a,b). In order to achieve complete oxidation, oxygenenriched air may be blown in during roasting. Sixty percent by weightof Na2CO3 was used without an oxidation aid to obtain recoveries of96–98% of molybdenum and vanadium after roasting for 1–3 h. Themolybdenum and vanadium salts were then leached with water in atemperature range from 60 °C to the boiling point of the solution.

A direct sodium carbonate roasting/water leaching process wasreported (Parkinson and Ishio, 1987), with ammonium chloride beingadded to precipitate ammoniummetavanadate. The solutionwas thenheated to 80–85 °C and acidified to precipitate molybdic acid whichcan be calcined tomolybdenum trioxide. One problemwas that duringroasting the cobalt and nickel formed stable compounds with aluminawhich were difficult to dissolve. Smelting should be a possible routefor recovering the cobalt and nickel.

A process flow sheet involving a sodium carbonate roasting/waterleaching proposed by Sebenik and Ference (1982) is shown in Fig. 8.Recoveries of both molybdenum and vanadium reached over 90%.

2.5.3. Direct sodium carbonate leachingDirect sodium carbonate leaching is similar to the salt roasting in

that molybdenum and vanadium were converted to soluble sodiumsalt under high pressures of 54 to 170 atm to oxidise the sulphides andcarbon (Sebenik et al., 1985). At the same time, SO3 produced duringoxidation was neutralised to environmentally acceptable Na2SO4 bysodium carbonate. The resultant solution containing mixed metalswas then separated by conventional separation techniques such asprecipitation, solvent extraction and ion exchange. In this process, allof the sulphur was converted to sulphate unlike roasting where at

Fig. 8. A conceptual process flow sheet involving sodium carbonate roasting/

least some sulphur was converted to SO2, resulting in less gasemission to pollute the environment. The disadvantage of directsodium carbonate leaching is that high pressures are needed.

Park et al. (2006) proposed a process in which the leaching ofspent HDS catalysts with sodium carbonate and hydrogen peroxidewas followed by separation and recovery of molybdenum asmolybdenum trioxide from the leach solution by carbon adsorption/desorption and selective precipitation. Under optimum conditions, aMoO3 product with a purity of 99.4% was achieved.

2.6. Bioleaching

Biotechnological leaching is now sought as an alternative to allother leaching methods to recycle spent catalysts in order to avoidhigh costs and some negative environmental impact of conventionalmethods such as acid leaching and caustic leaching which maygenerate large volumes of potentially hazardous wastes and gaseousemissions (Mishra et al., 2007). Recycling of metals fromHDS catalystsusing biotechnology is still a commercially sensitive area and fewpapers can be found.

Molybdenum, vanadium and nickel have stable soluble species inthe aqueous solution at 25 °C, pHb2.0 and EhN500 mV. These are theright living conditions for the organisms from the chemo-autotrophic,acidophilic and sulphur oxidising genus Acidithiobacillus (Pourbaix,1974). In such an environment, sulphuric acid is generated throughthe biooxidation of sulphur and/or reduced-sulphur compounds byacidophiles. Staley et al. (1989) reported that Acidithiobacillusproduces several intermediate sulphur species such as thiosulphateand sulphite with significant reducing properties while growing onelemental sulphur. The formation of sulphite can be expressed in thefollowing reaction:

S þ O2 þ H2O ¼ SO2−3 þ 2H

þ: ð7Þ

The sulphite can then be oxidised by oxygen and catalysed by thebacteria which thrive on low pH and tolerate harsh conditions inconcentrated metal solutions. A. thiooxidans and A. ferrooxidans werereported to reduce vanadium (V) to vanadium (IV) in the presence ofelemental sulphur (Briand et al., 1996; Bredberg et al., 2004).

Mishra et al. (2007) studied the bioleaching of spent refinerycatalysts using Acidithiobacillus type of bacteria. A one-step bioleach-ing process and a two-step bioleaching process were compared. Theone-step process involved inoculation of bacteria with solid spentcatalysts and elemental sulphur while the two-step process wasperformed by adding solid spent catalysts to previously culturedbacterial medium. The difference between the two processes was that

water leaching of spent catalysts (based on Sebenik and Ference, 1982).

8 L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

the biologically produced acid was the major leaching reagent in thetwo-step process where there was no direct contact between themetals in spent catalysts and the biomass. Under optimum conditions,Mo, V and Ni in the pre-treated spent catalyst were dissolved moreefficiently in the two-step rather than in the one-step bioleachingprocess because higher acid concentrations could be generated in thetwo-step process. Compared with direct chemical leaching withsimilar concentrations of sulphuric acid, bioleaching gives a slightlyhigher metal recovery.

Bio-hydrometallurgical processes are cost-effective, detoxify theresidues and reduce pollution of the environment (Brombacher et al.,1997). Aung and Ting (2005) also suggested the bio-hydrometallur-gical approach appears to offer good prospects for recovering valuablemetals from spent HDS in the future although much more researchwork is needed for commercialisation.

3. Summary

Among all the metallurgical processes for spent HDS catalysts, aprocess consisting of roasting and sulphuric acid leaching seems to bethe best option since all of the valuable metals dissolve. However, itconsumes large amounts of acid since the alumina carrier, the bulk ofthe catalyst, also dissolves and the downstream processes to produceseparate products of high purity are relatively complex. Sodiumcarbonate roasting followed by water leaching is a good option sincemolybdenum and vanadium are selectively extracted over aluminium,nickel and cobalt. Bioleaching offers good prospects for recoveringvaluable metals and at the same time generates much less environ-mental pollution. However, muchmore research work is needed for itscommercialisation. In all the leaching routes, the metals in thesolution have to be separated and purified by conventional separationtechniques such as precipitation, solvent extraction, ion exchange andadsorption methods. These are the focus of Part II of this review.

Acknowledgments

The authors would like to thankMs. Sue Cook, Dr. Xianwen Dai andMr. Yoko Pranolo for collecting some reference papers, Dr. WenshengZhang and Dr. Matthew Jeffrey for reviewing this paper and providingvaluable comments.

References

Anon, 1992. Recovery of metals from spent catalysts. General Information Leaflet ofMETREX.

Anon, 2003. Vanadium Statistics and Information in Mineral Commodity Summaries,2003. U.S.A. Government Minerals Publications Website.

Aung, K.K.M., Ting, Y.P., 2005. Bioleaching of spent fluid catalytic cracking catalyst usingAspergillus niger. Journal of Biotechnology 116, 159–170.

Biswas, R.K., Wakihara, M., Taniguchi, M.,1985. Recovery of vanadium andmolybdenumfrom heavy oil desulphurization waste catalyst. Hydrometallurgy 14, 219–230.

Biswas, R.K., Wakihara, M., Taniguchi, M., 1986. Characterisation and leaching of theheavy oil desulphurisation waste catalyst. Bangladesh Journal of Scientific andIndustrial Research 21, 228–237.

Bredberg, K., Karlsson, H.T., Holst, O., 2004. Reduction of vanadium (V) with Aci-dithiobacillus ferrooxidans and Acidithiobacillus thiooxidans. Bioresource Technology92, 93–96.

Briand, L., Thomas, H., Donati, E., 1996. Vanadium (V) reduction in Thiobacillusthiooxidans cultures on elemental sulfur. Biotechnology Letters 18, 505–508.

Brombacher, C., Bachofen, R., Brandl, H., 1997. Biohydrometallurgical processing ofsolids: a patent view. Applied Microbiology and Biotechnology 48, 577–587.

Chen, Y., Feng, Q.M., Shao, Y.H., Zhang, G.F., Ou, L.M., Lu, Y.P., 2006. Investigations on theextraction of molybdenum and vanadium from ammonia leaching residue of spentcatalyst. International Journal of Mineral Processing 79, 42–48.

Crnojevich, R.,Wiewiorowski, E., Tinnin, L., Case, A.,1990. Recycling chromium–aluminiumwastes from aluminium finishing operations. Journal of Metals 42 (10), 42–45.

Furimsky, E., 1996. Spent refinery catalysts: environment, safety and utilization.Catalysis Today 30, 223–286.

Gaballah, I., Djona, M., 1994. Valuable metals recovery from spent catalysts by selectivechlorination. Resources, Conservation and Recycling 10, 87–96.

Gaballah, I., Djona, M., 1995. Recovery of Co, Ni, Mo and V from unroasted spenthydrorefining catalysts by selective chlorination. Metallurgical and MaterialsTransactions B 26B, 41.

Grzechowiak, J., Grysiewicz, W., Ostrowski, A., Radomyski, B., Walendziewski, J., 1987.Recovery of aluminium, cobalt, iron and nickel from used catalysts containingmetals of ferrous group and possibly molybdenum on alumina support. Polish Pat.No. 136713, 1987.

Ho, E.M., 1992. Recovery of metals from spent catalysts. Honours Thesis, MurdochUniversity, Perth, Australia.

Howard, R.A., Barnes, W.R., 1991. Smelting process for recovery of valuable metals fromspent catalysts on an oxide support. US Pat. No. 5013533, 1991.

Hyatt, D.E., 1987. Value recovery from spent alumina-base catalyst. US Pat. No. 4657745,1987.

Ikeyama, T., 1987. Recovery of high purity vanadium and molybdenum componentsfrom oxidation-calcined spent catalysts. Japanese Pat. No. 62102834, 1987.

Jocker, J.S.M., 1993. The Metrex process, full recycling of spent hydroprocessingcatalysts. Symposium on Regeneration, Reactivation and Reworking of SpentCatalysts, 205th National Meeting, American Chemical Society, Denver, CO, March28–April 2, pp. 77–80.

Jong, B.W., Siemens, R.E., 1985. Proposed methods for recovering critical metals fromspent catalysts. Recycle and Secondary Recovery of Metals, Proc. of the Interna-tional Symposium, pp. 477–488.

Jong, B.W., Rhoads, S.C., 1989. Recovery of Principal Metal Values from WasteHydroprocessing Catalysts. Bureau of Mines, US Department of Interior. RI 9252.

Kar, B.B., Datta, P., Misra, V.N., 2004. Spent catalyst: secondary source for molybdenumrecovery. Hydrometallurgy 72, 87–92.

Kar, B.B., Murthy, B.V.R., Misra, V.N., 2005. Extraction of molybdenum from spentcatalyst by salt-roasting. International Journal of Mineral Processing 76, 143–147.

Krismer, B., Mueller, H.R., Nadler, H.G., 1979. Recycling of the valuable componentscobalt and molybdenum from hydrodesulphurisation catalysis in consideration ofmethods conserving the environment. Erzmetall 32 (12), 514–518.

Lee, F.M., Knudsen, R.D., Kidd, D.R., 1992. Reforming catalyst made from the metalsrecovered from spent atmospheric residue of desulphurisation catalyst. Industrial &Engineering Chemistry Research 31, 487–490.

Medvedev, A.S., Malochkina, N.V., 2007. Sublimation of molybdenum trioxide fromexhausted catalysts employed fro the purification of oil products. Metallurgy ofRare and Noble Metals 2, 31–34.

Mihashi, T., Motomura, H., Takeuchi, H., 1982. Recovering processes of valuable metalsfrom spent petroleum hydrodesulphurisation catalysts. Japanese Pat. No. 57022119,1982.

Millsap, W.A., Reisler, N., 1978. Cotter's new plant diets on spent catalysts … andrecovers molybdenum, nickel, tungsten and vanadium products. Engineering andMining Journal 179 (5), 105–107.

Mishra,D., Kim,D.J., Ralph,D.E., Ahn, J.G., Rhee,Y.H., 2007.Bioleachingof vanadiumrich spentrefinery catalysts using sulphur oxidising lithotrophs. Hydrometallurgy 88, 202–209.

Mueller, H.R., Krismer, B., Scottman, W., 1980. Recovery of valuable metals fromcatalysts. West Germany Pat. No. 2908570, 1980.

Naka, I., 1972. Recovery of vanadium and molybdenum from vanadium-containingmolybdenum spent catalysts. Japanese Pat. No. 47031892, 1972.

Ogui, N.K., Travkin, N.S., Lastovitskaya, K.S., Zaiko, V.P., Ryss, M.A., 1971. USSR Pat.No. 319337, 1971.

Oshiumi, T., 1979. Recovery of molybdenum and cobalt from a spent molybdenum–

cobalt–aluminium oxide catalyst. Japanese Pat. No. 54010215, 1979.Park, K.H., Mohapatra, D., 2007. Two stage leaching of activated spent HDS catalyst and

solvent extraction of aluminium using organo-phosphinic extractant, Cyanex 272.Journal of Hazardous Materials 148, 287–295.

Park, K.H., Mohapatra, D., Reddy, B.R., 2006. Selective recovery of molybdenum fromspent HDS catalyst using oxidative soda ash leach/carbon adsorption method.Journal of Hazardous Materials 138, 311–316.

Parkinson, G., Ishio, S., 1987. Recyclers try newways to process spent catalysts. ChemicalEngineering 94, 25–31.

Pourbaix, M., 1974. Atlas of Electrochemical Equilibria in Aqueous Solutions. NationalAssociation of Corrosion Engineers, Houston, Texas.

Rastas, J.K., Karpale, K.J., Tiitinen, H., 1983. Recovery of metal values from spent catalystsused in extracting sulphur from crude petroleum. Belgian Pat. No. 894678, 1983.

Sebenik, R.F., Ference, R.A., 1982. Recovery of metal values from spent cobalt–molybdenum/alumina petroleum hydrodesulphurisation and coal liquefactioncatalysts: laboratory-scale process and preliminary economics. Preprints —

American Chemical Society. Division of Petroleum Chemistry 27 (3), 674–678.Sebenik, R.F., Lavalle, P.P., Laferty, J.M., May, W.A., 1985. Recovery of metal values from

spent hydrodesulphurisation catalysts. US Pat. No. 4495157, 1985.Sefton, V.B., Fox, R., Lorenz, W.P., 1989. Recovery and separation of metals from spent

petroleum refining catalyst. US Pat. No. 4861565, 1989.Siemens, R.E., Jong, B.W., Russell, J.H., 1986. Potential of spent catalysts as a source of

critical metals. Conservation and Recycling 9 (2), 89–96.Staley, J.T., Bryant, M.P., Pfenning, N., Holt, J.F. (Eds.), 1989. Bergey's Manual of

Systematic Bacteriology, vol. 3. Williams and Wilkins, USA. 1857.Sun, D., Li, X.Z., Brungs, M., Trimm, D., 1998. Encapsulation of heavy metals on spent

fluid catalytic cracking catalyst. Water Science and Technology 38, 211–217.Sun, D., Tay, J.H., Cheong, H.K., Leung, D.L.K., Qian, G.R., 2001. Recovery of heavy metals

and stabilization of spent hydrotreating catalyst using a glass–ceramic matrix.Journal of Hazardous Materials B 87, 213–223.

Suzuki and Gao, S.L., 1982. Recovery of valuable metals in spent heavy oilhydrodesulphurisation catalyst. Japanese Pat. No. 57082122, 1982.

Toda, S.,1989a. Alkaline roastingof spent catalyst containingmolybdenumand vanadium.Journal of the Mining and Materials Processing Institute of Japan 105 (3), 261–264.

Toda, S., 1989b. Recovery of molybdenum and vanadium from spent desulphurisingcatalyst. Journal of the Mining and Materials Processing Institute of Japan 105 (3),265–269.

9L. Zeng, C.Y. Cheng / Hydrometallurgy 98 (2009) 1–9

Toda, S., Morimoto, M., 1983. Apparatus for roasting of spent catalyst. Japanese Pat. No.58020716, 1983.

Toda, S., Ishikawa, H., Yoshida, Y., 1979. Oxidation calcination of waste catalyst forpetroleum refining. Japanese Pat. No. 54107495, 1979.

Van den Berg, J.A.M., Yang, Y., Nauta, H.H.K., van Sandwijk, A., Reuter, M.A., 2002.Comprehensive processing of low grade sulphidic molybdenum ores. MineralsEngineering 15, 879–883.

Villarreal, M.S., Kharisov, B.I., Torres-Martinez, L.M., Elizondo, V.N., 1999. Recovery ofvanadium and molybdenum from spent petroleum catalyst of PEMEX. Industrial &Engineering Chemistry Research 38, 4624–4628.

Ward, V.C., 1989. Meeting environmental standards when recoveringmetals from spentcatalyst. Journal of Metals 41 (1), 54–55.

Wiewiorowsk, E.I., Tinnin, L.R., Crnojevich, R., 1987. Cyclic process for recovering metalvalues and alumina from spent catalysts. US Pat. No. 4670229, 1987.

Yasuhara, H., 1982. Treatment of waste desulphurisation catalyst. Japanese Pat.No. 57003716, 1982.

Zhang, Q.X., Zhao, Q.S., 2005. Metallurgy of Tungsten andMolybdenum. Metallurgy andIndustry Publisher, Beijing, China. (in Chinese).

Zhang, P.W., Inoue, K., Yoshizuka, K., Tsuyama, H., 1995. Recovery of metal values fromspent hydrodesulfurisation catalyst by solvent extraction with PIA-8. JapanChemistry 5, 412 (in Japanese).

Zhang, P.W., Inoue, K., Yoshizuka, K., Tsuyama, H., 1996. Extraction and selectivestripping of molybdenum (VI) and vanadium (IV) from sulphuric acid solutioncontaining aluminium (III), cobalt (II), nickel (II) and iron (III) by LIX 63 in ExxsolD80. Hydrometallurgy 41, 45–53.