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Article Separation of Rhenium from -Rich Concentrate via Hydrochloric Leaching Followed by Oxidative Roasting

Guanghui Li, Zhixiong You *,†, Hu Sun, Rong Sun, Zhiwei Peng *,†, Yuanbo Zhang and Tao Jiang

School of Processing & Bioengineering, Central South University, Changsha 410083, China; [email protected] (G.L.); [email protected] (H.S.); [email protected] (R.S.); [email protected] (Y.Z.); [email protected] (T.J.) * Correspondence: [email protected] (Z.Y.); [email protected] (Z.P.); Tel.: +86-731-888-30542 (Z.Y. & Z.P.) † These authors contributed equally to this work.

Academic Editor: Hugo F. Lopez Received: 12 October 2016; Accepted: 13 November 2016; Published: 16 November 2016

Abstract: Lead-rich molybdenite is a typical rhenium-bearing resource in China, which has not been efficiently utilized due to its high contents of lead and gangue minerals. In this study, was used for preliminarily removing lead and calcite from a lead-rich molybdenite concentrate. Oxidative roasting-ammonia leaching was then carried out for separation of rhenium and extraction of molybdenum. The hydrochloric acid leaching experiments revealed that 93.6% Pb and 97.4% Ca were removed when the leaching was performed at 95 ◦C for 10 min with HCl concentration of 8 wt. % and liquid- ratio of 5 (mL/g). The results of direct oxidative roasting indicated that 89.3% rhenium was volatilized from the raw concentrate after roasting at 600 ◦C for 120 min in air. In contrast, the rhenium volatilization was enhanced distinctly to 98.0% after the acid-leached concentrate (leaching residue) was roasted at 550 ◦C for 100 min. By the subsequent ammonia leaching, 91.5% molybdenum was leached out from the calcine produced from oxidative roasting of the acid-leached concentrate, while only 79.3% Mo was leached from the calcine produced by roasting molybdenite concentrate without pretreatment.

Keywords: molybdenite concentrate; rhenium; lead; oxidative roasting; acid leaching

1. Introduction Rhenium (Re) is an important that has been widely used in many fields, such as national defense, petrochemical engineering, aerospace, photovoltaics, solar cells, etc [1]. The global Re resource reached 2.5 million kg with only 48,800 kg produced in 2014. More than half of the production was from Chile, which has more than 1.3 million kg Re in reserves [2]. Rhenium is a rare metal occurring as (ReS2) mainly accompanying molybdenite. It was also found in some , niobium yttrium iron, and uranium ores. Typically, copper or molybdenum sulfide concentrates via flotation circuits contain 50–100 g/t (ppm) Re [3]. Because of the high value of rhenium, recovering rhenium from molybdenum concentrate has received much attention. During the roasting process of molybdenite concentrate, rhenium is volatized and enters the gas as Re2O7. A small portion may remain with the calcine. There are two types of roasting: oxidative roasting and solidification roasting [4]. The former process recovers molybdenum by leaching the calcine in ammonium hydroxide to produce ammonium molybdate. The contained rhenium is separated and volatilized into dust in the form of (Re2O7), which are soluble in water or peroxide solutions [5]. In solidification roasting, both rhenium and molybdenum

Metals 2016, 6, 282; doi:10.3390/met6110282 www.mdpi.com/journal/metals Metals 2016, 6, 282 2 of 12 are retained in calcine in the presence of slaked lime [6] or soda ash [7]. Rhenium is then extracted by leaching the lime-roasted calcine in or water [6]. A common problem with the solidification roasting method is attributed to the large amount of slaked lime added, which would lower the contents of molybdenum and rhenium and prolong the enrichment , exerting an unfavorable effect on the recovery of rhenium. Unlike lime-roasting, the molybdenum and rhenium in soda ash-roasted calcine could be leached in water solution [8]. Molybdenum and rhenium can be simultaneously leached by pressure oxidation leaching (POX), which is divided into two types: acid pressure leaching [9–11] and alkaline pressure leaching according to the leaching medium [12]. The use of high pressure and high concentration acid or alkali corroded the equipment severely. These two metals can also be extracted by other approaches, such as electrochemical leaching [13,14], bioleaching [15,16] and microwave-assisted leaching [17], although they have not been utilized commercially. The current industrial practice in the recovery of rhenium from rhenium-bearing solutions was generally conducted through ion exchange or solvent extraction [18–21]. Up to now, oxidative roasting has widely been used around the world by large manufacturers because of its low cost, easy operation and high efficiency [22,23]. However, the formation of molybdate and sulfate due to the existence of sulfides (PbS, FeS, ZnS and Cu2S) often restrains the recovery of molybdenum [24]. Moreover, the reactions between gangue minerals and molybdite would result in serious , which restricts the oxidation of molybdenite and volatilization of rhenium [25]. Previous studies mainly focused on the behavior of molybdenum during the roasting [26,27], but few studies explored the behaviors of rhenium and the gangue minerals in favor of separating rhenium from molybdenum. The volatilization behavior of rhenium and roasting characteristics of molybdenum concentrate deserve further attention. Lead-rich molybdenite is a typical molybdenum resource in Shanxi province of China, which has not been efficiently utilized mainly due to its high lead content [28]. The successful roast–leach process should volatilize rhenium as much as possible but simultaneously reduce the volatilization of molybdenum, leaving a negligible effect on the subsequent ammonium leaching process. In view of the high lead content in the raw material, hydrochloric acid was adopted to initially remove lead and calcite in this study [29]. The effects of leaching parameters on the removal of lead and the separation of rhenium during subsequent oxidative roasting were examined. Eventually, molybdenum was leached from the calcine in the ammonium solution to identify the favorable effect of HCl leaching.

2. Experimental Section

2.1. Materials The lead-rich molybdenite concentrate sample was obtained from a dressing plant of Shanxi province, China. The size distribution of the sample was 95.2 wt. % of the particles smaller than 0.074 mm. The main chemical composition of the sample is shown in Table1. The Mo and Re contents were 43.55 wt. % and 321 g/t, respectively, and the lead content was as high as 4.52 wt. %. The XRD pattern shown in Figure1 indicates that molybdenum mainly existed in the form of molybdenite (MoS 2) and the gangue minerals were galena (PbS), quartz (SiO2) and a small amount of calcite (CaCO3). The gases for oxidative roasting and cooling were mixed O2/N2 and N2 gases, respectively. The purity of these gases was above 99.99%.

Table 1. Main chemical composition of lead-rich molybdenite concentrate/wt. %.

Mo Re Fe SiO2 Al2O3 CaO MgO WO3 Cu Pb S 43.55 321 g/t 1.39 4.02 0.18 3.27 0.48 1.66 0.03 4.52 32.46 Metals 2016, 6, 282 3 of 12 Metals 2016, 6, 282 3 of 12

M Metals 2016, 6, 282 16000 3 of 12 M-Molybdenite, MoS2 G-Galena, PbS 15000 Q-Quartz, SiO M 2 16000 M-Molybdenite, MoS C-Calcite, CaCO3 2 14000 G-Galena, PbS 15000 Q-Quartz, SiO2

C-Calcite, CaCO3

1500 14000 M

Intensity/cps M 1000 M

1500 M M 500 Intensity/cps M M M 1000 G C M M MM M Q CQC G G C G M M MMM M 0 M 500 10 20 M30M 40 50 60 70 80 G C M MM o M Q CQC G Two-theta/G C G M M MMM M 0 Figure 1. XRD10 pattern20 of the30 lead-rich40 molybdenum50 60 70 concentrate.80 Figure 1. XRD pattern of the lead-rich molybdenumo concentrate. Two-theta/ 2.2. Experimental2.2. Experimental Procedure Procedure Figure 1. XRD pattern of the lead-rich molybdenum concentrate. TheThe experimental experimental procedure procedure mainly mainly includedincluded acid acid leaching, leaching, drying, drying, oxidative oxidative roasting, roasting, cooling, cooling, etc.2.2. Inetc. eachExperimental In each leaching leaching Procedure test, test, 5 5 g g of of the the sample sample was treated treated in in a a beaker beaker containing containing a predetermined a predetermined quantity of hydrochloric acid (determined according to the liquid-solid ratio (mL/g). The beaker was quantity ofThe hydrochloric experimental acidprocedure (determined mainly included according acid to leaching, the liquid-solid drying, oxidative ratio (mL/g). roasting, The cooling, beaker was located in a water bath that maintains the leaching temperature with a solution stirred continuously. locatedetc. in In a each water leaching bath that test, maintains 5 g of the the sample leaching was treated temperature in a beaker with containing a solution a stirred predetermined continuously. After leaching for a desired duration, the solution was filtered and the residue obtained was washed Afterquantity leaching of forhydrochlor a desiredic acid duration, (determined the solution according was to the filtered liquid and-solid the ratio residue (mL/g) obtained. The beaker was was washed with hot distilled water and dried for chemical analysis. withlocated hot distilled in a water water bath and that dried maintains for chemical the leaching analysis. temperature with a solution stirred continuously. The samples after hydrochloric acid leaching (acid-leached concentrate) were balled into green After leaching for a desired duration, the solution was filtered and the residue obtained was washed Thepellets samples with 10– after14 mm hydrochloric in diameter. The acid pellets leaching were (acid-leached dried in an oven concentrate) at 80 °C for 4 were h before balled oxidative into green with hot distilled water and dried for chemical analysis. ◦ pelletsroasting. with 10–14 The schematic mm in diameter. diagram Theof the pellets roasting were apparatus dried in is an shown oven in at Figure 80 C 2. for The 4 h dried before pellets oxidative The samples after hydrochloric acid leaching (acid-leached concentrate) were balled into green roasting.were Thefirstly schematic loaded into diagram the reactor of the (a roasting quartz tube apparatus with perforated is shown screen in Figure plate2. at The the dried bottom), pellets into were pellets with 10–14 mm in diameter. The pellets were dried in an oven at 80 °C for 4 h before oxidative firstlywhich loaded the into nitrogen the reactor gas was (a flushed quartz to tube expel with the perforated air. Then, the screen quartz plate tube at was the bottom),transferred into into which the the roasting. The schematic diagram of the roasting apparatus is shown in Figure 2. The dried pellets heating zone of an electric resistance furnace that was preheated at a rate of 10 °C/min to a required nitrogenwere gasfirstly was loaded flushed into to the expel reactor the (a air. quartz Then, tube the with quartz perforated tube was screen transferred plate at intothe bottom), the heating into zone temperature. The mixed O2/N2 gases were directed into the tube immediately.◦ The composition of of anwhich electric the resistance nitrogen gas furnace was flushed that was to expel preheated the air. at Then a rate, the of quartz 10 C/min tube was to atransferred required temperature.into the oxidative gas was adjusted by the flowmeters of O2 and N2, and the total flow rate was fixed at 4.0 The mixedheating Ozone2/N of2 gasesan electric were resistance directed furnace into the that tube was immediately. preheated at a Therate compositionof 10 °C/min to of a oxidativerequired gas L/min to ensure reaction atmosphere. After roasting for a certain time period, the quartz tube was was adjustedtemperature. by theThe flowmeters mixed O2/N of2 gases O2 and were N 2directed, and the into total the flow tube rateimmediately. was fixed The at 4.0composition L/min to of ensure taken out of the furnace and cooled down in N2. The calcine was then collected for analysis. reactionoxidative atmosphere. gas was adjusted After roasting by the flowmeters for a certain of O time2 and period, N2, and thethe quartztotal flow tube rate was wastaken fixed at out 4.0 of the L/min to ensure reaction atmosphere. After roasting for a certain time period, the quartz tube was furnace and cooled down in N2. The calcine was then collected for analysis. taken out of the furnace and cooled down in N2. The calcine was then collected for analysis.

Figure 2. Schematic diagram of the equipment for roasting.

Metals 2016, 6, 282 4 of 12

2.3. Analytical Method The XRD analysis was conducted using a diffractometer (RIGAKU D/Max 2500, Akishima, Japan) under the conditions as follows: Cu Kα, tube current and voltage: 250 mA, 40 kV, scanning range: 10◦–80◦ (2θ), step size: 0.02◦ (2θ) and scanning speed: 8◦/min. The contents of Pb and Mo were measured by the chemical titration method, while the Re content was assayed by the instrument of Inductively Coupled Plasma Optical Emission Spectrometer (ICP-OES, Thermo Scientific iCAP 7000 series, Waltham, MA, USA). The acidic leaching ratio of Pb, Mo or Re was calculated by the equation as follows:

m × ε β = (1 − 1 1 ) × 100% (1) m0 × ε0 where β is the leaching ratio, %; m0 is the weight of molybdenite concentrate, g; ε0 is the weight fraction of Pb, Mo or Re of molybdenite concentrate, %; m1 is the weight of leached residue, g; and ε1 is the Pb, Mo or Re weight fraction of leached residue, %. The volatilization ratio of Mo or Re was calculated according to the following equation:

m × ε α = (1 − 2 2 ) × 100% (2) m0 × ε0 where α is the volatilization ratio (%); m0 is the weight of molybdenite concentrate, g; ε0 is the Mo or Re weight fraction of molybdenite concentrate, %; m2 is the weight of roasted product, g; and ε2 is the weight fraction of Mo or Re of roasted product, %.

3. Results and Discussion

3.1. Removal of Pb by Acid Leaching with HCl

3.1.1. Chemistry of HCl Leaching In HCl solution, galena is capable of being selectively removed due to its high solubility. Conversely, molybdenite remains chemically stable. Traditional processes of removing galena from molybdenite concentrate with HCl are carried out in the presence of CaCl2 or NaCl [28]. Due to the existence of CaCO3 (Table1) in the raw concentrate, the newly formed CaCl 2 may facilitate the dissolution of PbCl2. Thus, no additional CaCl2 is needed during acid leaching. The chemical reactions can be described by the following equations:

CaCO3 + 2HCl → CaCl2 + CO2 + H2O (3)

PbS + 2HCl → PbCl2 + H2S (4) − 2− PbCl2 + 2Cl → [PbCl4] (5) During the acid leaching, calcite is easily dissolved (Equation (3)) and galena will react with hydrochloric acid to form lead chloride (Equation (4)). The lead chloride is soluble in hot water, and its solubility is improved distinctly in CaCl2 solution by forming a complex compound (Equation (5)). As a result, the galena and calcite can be effectively removed and lead chloride is expected to be recovered by cooling the leachate [30].

3.1.2. Effect of HCl Concentration The effect of hydrochloric acid concentration on the removal of Pb was investigated by keeping leaching temperature at 95 ◦C, leaching time of 10 min and liquid-solid ratio of 5 (mL/g). The results are plotted in Figure3. It can be observed that lead was not effectively removed until the hydrochloric acid concentration was higher than 6 wt. %. By increasing the HCl concentration from 1 wt. % to MetalsMetals 20162016,, 66,, 282282 55 ofof 1212 areareMetals plotted plotted2016, 6, 282 in in Figure Figure 3. 3. It It can can be be observed observed that that lead lead was was notnot effectively effectively removed removed until until5 of the the 12 hydrochlorichydrochloric acidacid concentrationconcentration waswas higherhigher thanthan 66 wt.wt. %%.. ByBy increasingincreasing thethe HClHCl concentrationconcentration fromfrom 11 wt. % to 6 wt. %, the removal of Pb increased dramatically from 23.4% to 92.5%, and then it remained wt.6 wt. % %,to 6 the wt. removal %, the removal of Pb increaseddramatically of Pb increased dramatically from 23.4% from to 23.4% 92.5%, to and 92.5%, then and it remainedthen it remained almost almost constant. During the acid leaching, hydrochloric acid is not only used to remove galena, but almostconstant. constant. During Duri theng acid the leaching, acid leaching, hydrochloric hydrochloric acid isacid not is only not usedonly used to remove to remove galena, galena, but alsobut also exhausted by CaCO3, MgCO3 as well as other . The HCl concentration required for alsoexhausted exhausted by CaCO by CaCO, MgCO3, MgCOas well3 as as well other as sulfides. other sulfides. The HCl The concentration HCl concentration required required for leaching for leaching galena is actually3 a3 small part. The stoichiometric threshold of HCl concentration for leachinggalena is galenaactually is a actuallysmall part. a small The stoichiometric part. The stoichiometric threshold of threshold HCl concentration of HCl concen for removingtration the for removing the majority of Pb, Ca, Mg and part of Fe is about 1.5%–2%. As observed from Figure 3, the removingmajority of the Pb, majority Ca, Mg of and Pb, partCa, Mg of Feand is part about of 1.5%–2%.Fe is about As 1.5% observed–2%. As fromobserved Figure from3, the Figure dramatic 3, the dramatic increase in Pb extraction is 2–3 times the stoichiometric threshold. Meanwhile, the leaching dramaticincrease inincrease Pb extraction in Pb extraction is 2–3 times is 2–3 the times stoichiometric the stoichiometric threshold. threshold. Meanwhile, Meanwhile, the leaching the leaching ratio ratio of molybdenum increased gradually to its maximum value of 1.7% as the HCl concentration ratioof molybdenum of molybdenum increased increased gradually gradually to its to maximum its maximum value value of 1.7% of 1.7% as the as HCl the HCl concentration concentration was was increased to 6 wt. %. A further increase in HCl concentration lowers the leaching ratio of wasincreased increased to 6 wt. to %. 6 wt. A further %. A increase further in increase HCl concentration in HCl concentration lowers the leaching lowers the ratio leaching of molybdenum. ratio of molybdenum. In contrast, the leaching ratio of rhenium kept increasing with HCl concentration. The molybdenum.In contrast, the In leaching contrast, ratio the leaching of rhenium ratio kept of rhenium increasing kept with increasing HCl concentration. with HCl concentration. The suitable HClThe suitable HCl concentration was 8 wt. %. suitableconcentration HCl concentration was 8 wt. %. was 8 wt. %.

1010 100100

/% 8 80 /% 8 80

/%

/%

66 6060 MoMo ReRe 44 4040 PbPb

Leaching of Pb

Leaching of Pb 2 20 Leaching of Mo and Re 2 20

Leaching of Mo and Re

00 00 00 55 1010 1515 2020 2525 3030 3535 4040 HCl concentration /wt.% HCl concentration /wt.% Figure 3. Effect of HCl concentration on the leaching of Mo, Re and Pb. FigureFigure 3.3. EffectEffect ofof HClHCl concentrationconcentration onon thethe leachingleaching ofof Mo,Mo, ReRe andand Pb.Pb. 3.1.3. Effect of Leaching Temperature 3.1.3.3.1.3. EffectEffect ofof LeachingLeaching TemperatureTemperature By fixing the HCl concentration at 8 wt. %, leaching time at 10 min and liquid-solid ratio at By fixing the HCl concentration at 8 wt. %, leaching time at 10 min and liquid-solid ratio at 5 5 (mL/g),By fixing the effectthe HCl of concentration leaching temperature at 8 wt. on%, theleaching removal time of at lead 10 min and and leaching liquid of-solid molybdenum ratio at 5 (mL/g), the effect of leaching temperature on the removal of lead and leaching of molybdenum and (mL/g),and rhenium the effect was of examined, leaching temperature as shown in on Figure the removal4. As theof lead leaching and leaching temperature of molybdenum was increased, and rheniumrhenium was was examined, examined, as as shown shown in in Figure Figure 4. 4. As As the the leaching leaching temperature temperature was was increased, increased, the the the reaction between galena and hydrochloric acid was accelerated and the solubility of PbCl2 was reaction between galena and hydrochloric acid was accelerated and the solubility of PbCl2 was reactionimproved. between The leaching galena ratios and hydrochloric of Pb and Re increasedacid was graduallyaccelerated from and 24.5% the solubility and 0.5% of to PbCl 93.6%2 was and improved. The leaching ratios of Pb and Re increased gradually from 24.5% and 0.5% to 93.6% and improved.3.6%, respectively, The leaching when ratios the leaching of Pb and temperature Re increased was gradually increased from from 24.5% 25 ◦ Cand to 950.5%◦C. to In 93.6% addition, and 3.6%, respectively, when the leaching temperature was increased from 25 °C to 95 °C. In addition, 3.6%,the little respectively, Mo was extracted when the in thisleaching regard, temperature and the appropriate was increased leaching from temperature 25 °C to 95 was°C. 95In addition,◦C. thethe littlelittle MoMo waswas extractextracteded inin thisthis regard,regard, andand thethe appropriateappropriate leachingleaching temperaturetemperature waswas 9595 °°C.C.

44 100100

/% Mo

/% Mo 80 3 80 3 ReRe

Pb /% Pb /% 6060 22 4040

Leaching of Pb 11 Leaching of Pb

Leaching of Mo and Re Leaching of Mo 20

Leaching of Mo and Re Leaching of Mo 20

00 00 2020 3030 4040 5050 6060 7070 8080 9090 100100 o Leaching temperature/ oC Leaching temperature/ C Figure 4. Effect of leaching temperature on the leaching of Mo, Re and Pb.

Metals 2016, 6, 282 6 of 12 Metals 2016, 6, 282 6 of 12 Metals 2016, 6, 282 6 of 12 Figure 4. Effect of leaching temperature on the leaching of Mo, Re and Pb. Figure 4. Effect of leaching temperature on the leaching of Mo, Re and Pb. 3.1.4.3.1.4. Effect Effect of of Leaching Leaching Time Time 3.1.4. Effect of Leaching Time TheThe effect effect of of leaching leaching time time on on the the removal removal of of Pb Pb was was studied studied under under the the following following conditions: conditions: The effect of leaching time on the removal of Pb was studied under the following conditions: HClHCl concentration concentration of of 8 8wt. wt. %, %, leaching leaching temperature temperature of 95 of 95°C ◦andC and liquid liquid-solid-solid ratio ratio of 5 of(mL/g). 5 (mL/g). The HCl concentration of 8 wt. %, leaching temperature of 95 °C and liquid-solid ratio of 5 (mL/g). The resultsThe results plotted plotted in Figure in Figure 5 indicate5 indicate that that 85.5% 85.5% of ofPb Pb was was leached leached in in only only 2 2 min min,, and and this numbernumber results plotted in Figure 5 indicate that 85.5% of Pb was leached in only 2 min, and this number increasedincreased gradually gradually to to 93.6% 93.6% in in 10 10 min. min. The The variation variation of of Re Re was was similar similar to to that that of of Pb, Pb, which which increased increased increased gradually to 93.6% in 10 min. The variation of Re was similar to that of Pb, which increased fromfrom 0.6% 0.6% to to 3.6% 3.6% within within the the time time range range of of 2 2–10–10 min. However, However, the the leaching leaching ratio ratio of of Mo Mo changed from 0.6% to 3.6% within the time range of 2–10 min. However, the leaching ratio of Mo changed slightlyslightly when when the the leaching leaching time time was was prolonged prolonged to 60 to min. 60 Hence,min. Hence, leaching leaching time of time 10 min of was 10 min sufficient was slightly when the leaching time was prolonged to 60 min. Hence, leaching time of 10 min was sufficientfor removing for removing the majority the ofmajority lead. of lead. sufficient for removing the majority of lead. 5 100 5 100

/%

/% 4 80 4 80

/%

3 60 /% 3 Mo 60 Mo Re Re 2 Pb 40 2 Pb 40

Leaching of Pb

Leaching of Pb Leaching of Mo and Re 1 20 Leaching of Mo and Re 1 20

0 0 00 10 20 30 40 50 60 0 0 10 20 30 40 50 60 Leaching time/min Leaching time/min FigureFigure 5. 5. EffectEffect of of leaching leaching time time on on the the leaching leaching of of Mo, Mo, Re Re and and Pb. Pb. Figure 5. Effect of leaching time on the leaching of Mo, Re and Pb. 3.1.5.3.1.5. Effect Effect of of Liquid Liquid-Solid-Solid Ratio Ratio 3.1.5. Effect of Liquid-Solid Ratio AimingAiming at at optimizing optimizing the liq liquid-soliduid-solid ratio (mL/g) for for complete complete removal removal of of Pb, Pb, the the experiments experiments Aiming at optimizing the liquid-solid ratio (mL/g) for complete removal of Pb, the experiments werewere performed with the liquid liquid-solid-solid ratio ranging from 2 to 6. The HCl concentration, leaching leaching were performed with the liquid-solid ratio ranging from 2 to 6. The HCl concentration, leaching temperaturetemperature an andd time time were were maintained maintained at at 8 8 wt. wt. % %,, 95 95 °◦CC and and 10 10 min, min, respectively. respectively. From From the the res resultsults temperature and time were maintained at 8 wt. %, 95 °C and 10 min, respectively. From the results shownshown in in Figure Figure 66,, itit is ascertained that the removal of Pb increased significantlysignificantly from 74.5% to more shown in Figure 6, it is ascertained that the removal of Pb increased significantly from 74.5% to more thanthan 90% 90% as as the the liquid liquid-solid-solid ratio ratio was was increased increased to 5. to The 5. The leaching leaching ratio ratio of Re of maintained Re maintained at less at than less than 90% as the liquid-solid ratio was increased to 5. The leaching ratio of Re maintained at less than 4%than when 4% when the liquid the liquid-solid-solid ratio ratiowas smaller was smaller than than5 and 5 increased and increased distinctly distinctly thereafter. thereafter. However, However, the 4% when the liquid-solid ratio was smaller than 5 and increased distinctly thereafter. However, the leachingthe leaching ratio ratio of Mo of Modecreased decreased by byincreasing increasing the the liquid liquid-solid-solid ratio ratio up up to to 5. 5. Thus, Thus, the the most most suitable suitable leaching ratio of Mo decreased by increasing the liquid-solid ratio up to 5. Thus, the most suitable liquidliquid-solid-solid ratio ratio was was 5. 5. liquid-solid ratio was 5.

8 100 8 100

/% 80 /% 6 Mo 80

6 Mo /%

Re /% Re Pb 60 Pb 60 4 4 40 40

Leaching of Pb

2 Leaching of Pb 2 20 Leaching of Mo and Re 20

Leaching of Mo and Re

0 0 0 2 3 4 5 6 0 2 3 4 5 6 Liquid-solid ratio/(mL/g) Liquid-solid ratio/(mL/g) Figure 6. Effect of liquid-solid ratio on the leaching of Mo, Re and Pb. Figure 6. Effect of liquid-solid ratio on the leaching of Mo, Re and Pb. Figure 6. Effect of liquid-solid ratio on the leaching of Mo, Re and Pb.

Metals 2016, 6, 282 7 of 12

Metals 2016, 6, 282 7 of 12 From the above analysis, the suitable leaching conditions for removing Pb are as follows: HCl concentrationFrom the above of 8 analysis, wt. %, leaching the suitable temperature leaching ofconditions 95 ◦C, leaching for removing time of Pb 10 are min as and follows: liquid-solid HCl ratioconcentration of 5. Most ofof galena8 wt. % (93.6%), leaching and temperature calcite (97.4%) of were95 °C, dissolved leaching intime the of hydrochloric 10 min and acid.liquid The-solid main chemicalratio of composition5. Most of gal ofena molybdenite (93.6%) and concentratecalcite (97.4%) after were the dissolved pretreatment in the is hydrochloric summarized acid. in Table The 2. It canmain be chemical observed composition that Mo and of Re molybdenite were enriched concentrate after leaching after due the to pretreatment the removal isof summarizedgalena and calcite. in Table 2. It can be observed that Mo and Re were enriched after leaching due to the removal of galena and calcite.Table 2. Main chemical composition of molybdenite concentrate after pretreatment.

ComponentTable Mo 2. Main Re chemical Fecomposition SiO2 ofAl molybdenite2O3 CaO concentrate MgO after WO pretreatment.3 Cu Pb S

Componentwt. % 50.25 Mo 363 g/tRe 1.07Fe 4.72SiO2 Al 0.532O3 CaO 0.10 MgO 0.18 WO 1.763 Cu 0.02 Pb0.34 S 33.50 wt. % 50.25 363 g/t 1.07 4.72 0.53 0.10 0.18 1.76 0.02 0.34 33.50 3.2. Separation of Rhenium by Oxidative Roasting 3.2. Separation of Rhenium by Oxidative Roasting Oxidative roasting followed by ammonia leaching is the most classical process for recovering Mo and ReOxidative from molybdenite roasting followed concentrate by ammonia [22,23]. During leaching the is oxidativethe most classical roasting, process molybdenum for recovering disulfide is oxidizedMo and Re to molybdenum from molybdenite trioxide, concentrate which is [22,23]. soluble During in the ammonia the oxidative solution, roasting, while molybdenum rhenium was ◦ volatilizeddisulfide into is oxidized the flue to dust molybdenum in the form trioxide of Re2O, which7 (whose is soluble vaporpressure in the ammonia is 1 atm at solution, 362 C). while In this section,rhenium oxidative was volatilized roasting into of the the concentrate flue dust in after the form acid of leaching Re2O7 (whose was conducted vapor pressure to investigate is 1 atm theat 362 effect of° theC). pretreatment In this section, on oxidativethe separation roasting of rhenium. of the concentrate after acid leaching was conducted to investigate the effect of the pretreatment on the separation of rhenium. 3.2.1. Effect of Roasting Temperature 3.2.1. Effect of Roasting Temperature The effect of roasting temperature on the separation of Re is shown in Figure7. Oxidative roastingThe was effect carried of roasting out for temperature 120 min under on the the separation air atmosphere. of Re is The shown results in Figure in Figure 7. Oxidative7 indicate thatroasting the volatilization was carried out of Refor increased120 min under slowly the and air atmosphere. then remained The almostresults in constant Figure 7 at indicate temperatures that the volatilization of Re increased slowly and then remained almost constant at temperatures above above 550 ◦C. Almost complete volatilization of rhenium was achieved for the concentrate after 550 °C. Almost complete volatilization of rhenium was achieved for the concentrate after pretreatment, 10% higher than that without pretreatment. This is mainly attributed to the generation of pretreatment, 10% higher than that without pretreatment. This is mainly attributed to the generation lead molybdate and calcium molybdate in the presence of galena and calcite, which results in serious of lead molybdate and calcium molybdate in the presence of galena and calcite, which results in sintering and restrains the diffusion of O2. Additionally, the formation of perrhenate restricted the Re serious sintering and restrains the diffusion of O2. Additionally, the formation of perrhenate volatilizationrestricted the [25 Re]. volatilization [25]. OnOn the the contrary, contrary, the the volatilization volatilization of of Mo Mo before before and and after after pretreatment pretreatment gradually gradually increased increased to to 5.3% ◦ and5.3% 6.2%, and respectively, 6.2%, respectively, when thewhen roasting the roasting temperature temperature was elevated was elevated to 600 to C.600 Moreover, °C. Moreover, a further a increasefurther inincrease roasting in temperatureroasting temperature had a significant had a significant impact onimpact the Moon the volatilization. Mo volatilization. This is This because is molybdenumbecause molybdenum oxides formed oxides during formed roasting during would roasting also would partially also sublimate partially even sublimate below even their below melting points.their Furthermore, melting points. higher Furthermore, temperature higher accelerated temperature their accelerated sublimation. their Considering sublimation. the Conside volatilizationring ofthe Re, volatilization the roasting of temperature Re, the roasting could temperature be decreased could from be decreased 600 ◦C to from 500–550 600 ◦°C forto 500 the–550 concentrate °C for afterthe pretreatment.concentrate after pretreatment.

10 100

/%

8 /% 80

6 60

4 40 Mo, unpretreated

Volatilization of Mo Mo, pretreated Volatilization of Re 2 Re, unpretreated 20 Re, pretreated 0 0 450 500 550 600 650 o Roasting temperature/ C Figure 7. Effect of roasting temperature on the volatilization of Mo and Re. Figure 7. Effect of roasting temperature on the volatilization of Mo and Re.

Metals 2016, 6, 282 8 of 12

3.2.2. Effect of Roasting Time The effect of roasting time on the separation of Re was investigated and the results are shown in Figure 8. The pellets were roasted at 600 °C (unpretreated) or 550 °C (pretreated) in air for different time periods from 10 min to 150 min. It is found that the Mo volatilization increased steadily with increasing roasting time, and the Metals 2016, 6, 282 8 of 12 volatilization ratio of Mo in the raw material was greater than that after pretreatment. This is mainly attributed to the higher requrement on the roasting temperature for the raw concentrate. 3.2.2.Similarly, Effect of Roastingthe volatilization Time ratio of Re increased gradually with the increase of roasting time from 10 min to 100 min and then varied slightly. It is also evident that the rhenium volatilization was The effect of roasting time on the separation of Re was investigated and the results are shown in always higher than that without pretreatment under the experimental conditions. The maximum Re Figure8. The pellets were roasted at 600 ◦C (unpretreated) or 550 ◦C (pretreated) in air for different volatilization ratio reached 98.0% in 100 min, which was greater than that (86.6%) for the untreated time periods from 10 min to 150 min. concentrate. Thus, 100 min is adequate for roasting the concentrate after hydrochloric acid leaching.

10 100

/% 8 80 /%

6 60

4 40 Mo, unpretreated

Volatilization of Mo Mo, pretreated Volatilization of Re 2 Re, unpretreated 20 Re, pretreated 0 0 0 20 40 60 80 100 120 140 160 Roasting time/min Figure 8. Effect of roasting time on the volatilization of Mo and Re. Figure 8. Effect of roasting time on the volatilization of Mo and Re.

It is found that the Mo volatilization increased steadily with increasing roasting time, and the 3.2.3. Effect of O2 Partial Pressure volatilization ratio of Mo in the raw material was greater than that after pretreatment. This is mainly attributedThe effect to the of higher O2 partial requrement pressure on (O the2/(O roasting2 + N2)) on temperature the separation for the of rawrhenium concentrate. from molybdenum is shownSimilarly, in Figure thevolatilization 9. The experiments ratio of were Re increased carried out gradually under the with conditions the increase of 600 of roasting °C, 120 timemin (unpretreated)from 10 min to and 100 min550 and°C, 100 then min varied (pretreated), slightly. It respectively. is also evident The that results the rhenium in Figure volatilization 9 indicate that was bothalways the higher Mo and than Re thatvolatilization without pretreatment ratios of the raw under concentrate the experimental reached conditions.their maximum The maximumvalues as the Re Ovolatilization2 partial pressure ratio was reached increased 98.0% to in 21%, 100 min, and whichthen decreased was greater gradually. than that The (86.6%) sintering for thephenomenon untreated becomesconcentrate. more Thus, serious 100 min under is adequate higher O for2 partial roasting pressure, the concentrate causing after more hydrochloric compact structures acid leaching. of the roasted pellets [25]. 3.2.3.After Effect the of O molybdenite2 Partial Pressure concentrate was pretreated by acid leaching, Mo volatilization kept increasing with increase in the O2 partial pressure. In contrast, the volatilization of Re was improved The effect of O2 partial pressure (O2/(O2 + N2)) on the separation of rhenium from molybdenum and then remained nearly stable when the O2 partial pressure exceeded 21%. The maximum Re is shown in Figure9. The experiments were carried out under the conditions of 600 ◦C, 120 min volatilization ratio of 98.0% was obtained at 21% of O2 partial pressure, while the corresponding Mo (unpretreated) and 550 ◦C, 100 min (pretreated), respectively. The results in Figure9 indicate that volatilization was only 4.5%. Thus, air atmosphere was recommended for oxidative roasting. both the Mo and Re volatilization ratios of the raw concentrate reached their maximum values as the O2 partial pressure was increased to 21%, and then decreased gradually. The sintering phenomenon becomes more serious under higher O2 partial pressure, causing more compact structures of the roasted pellets [25]. After the molybdenite concentrate was pretreated by acid leaching, Mo volatilization kept increasing with increase in the O2 partial pressure. In contrast, the volatilization of Re was improved and then remained nearly stable when the O2 partial pressure exceeded 21%. The maximum Re volatilization ratio of 98.0% was obtained at 21% of O2 partial pressure, while the corresponding Mo volatilization was only 4.5%. Thus, air atmosphere was recommended for oxidative roasting.

Metals 2016, 6, 282 9 of 12 Metals 2016, 6, 282 9 of 12

10 100

/%

8 /% 80

6 60

4 40 Mo, unpretreated

Volatilization of Re Volatilization of Mo Mo, pretreated 2 Re, unpretreated 20 Re, pretreated 0 0 15 20 25 30 O partial pressure/min 2

Figure 9. Effect of O2 partial pressure on the volatilization of Mo and Re. Figure 9. Effect of O2 partial pressure on the volatilization of Mo and Re. 3.2.4. Phase Transformation in the Roasting Process 3.2.4. Phase Transformation in the Roasting Process To identify the phase transformation of molybdenite concentrate during oxidative roasting, To identify the phase transformation of molybdenite concentrate during oxidative roasting, XRD analysis was performed for determination of the phase compositions of roasted products. MetalsXRD 2016analysis, 6, 282 was performed for determination of the phase compositions of roasted products.10 ofThe 12 The diffractograms of the raw material and roasted products are compared in Figure 10. diffractograms of the raw material and roasted products are compared in Figure 10.

As observed15000 in Figure 10, the diffraction patterns of roasted products are different from that of m Calcine of leached residue the raw molybdenite concentrate. Molybdenitem (MoS2) in the raw material disappeared, whereas molybdite (MoO100003) was formed according to the following equations [26,27]. Meanwhile, rheniite

(ReS2) was also oxidized5000 to rhenium (VII) (Rem2O7) whose vapor pressure reached 1 atm at 362.4 °C. Thus, rhenium oxide is absent inm Figurem 10, while molybdite was retained in the calcine during mm m m mm m 150000 oxidative roasting: m Calcine of raw material m 10000 2MoS2 + 7O2 → 2MoO3 + 4SO2 (6)

m 5000 4ReS2 + 15O2 →m 2Re2O7 + 8SO2 (7) m Wm WPmm m mm P Pm By roasting15000 the0 raw concentrate at 600 °C for 120 min under air atmosphere, galena (PbS) was M Hydrochloric acid leached residue oxidized to lead13500 oxide (PbO), which combined with molybdite to form wulfenite (PbMoO4). In the Intensity/cps M meantime, calcite (CaCO3) in the raw materialM was decomposed to (CaO). Similarly, 1000 M M M CaO would also react with molybdite to generateM powellite (CaMoOM4), as described by the following 500 M M equations: Q Q M M MM M MM 0 15000 M → Raw material 2PbS + 3O2 2PbOM + 2SO2 (8) M 1000 M PbO +C MoO3 → PbMoO4 (9) Q M 500 M M M3 → 2 M M Q CGCaCOC G M CaO + COC G M M M M MM (10) 0 5 10 15 20 25CaO30 + MoO35 3 40→ CaMoO45 50 4 55 60 65 70 75 80 (11) Two-theta/o As a result of hydrochloric acid leaching, the peaks of galena and calcite were disappeared. No obviousFigure diffraction 10. XRD peaks patterns of ofwulfenite the molybdenite and powellite concentrate except and those roasted of products.molybdite (M—Molybdenite, are observed in the Figure 10. XRD patterns of the molybdenite concentrate and roasted products. (M—Molybdenite, XRD MoSpattern2; G—Galena, of the roasted PbS; C—Calcite,product, which CaCO was3; Q—Quartz, obtained SiOat 5502; m—Molybdite, °C for 100 min MoO under3; P—Powellite, air atmosphere. MoSCaMoO2; G4—; W—Wulfenite,Galena, PbS; C PbMoO—Calcite,4). CaCO3; Q—Quartz, SiO2; m—Molybdite, MoO3; P—Powellite, CaMoO4; W—Wulfenite, PbMoO4). As observed in Figure 10, the diffraction patterns of roasted products are different from that of To evaluate the effect of removing lead and calcium on the leaching of Mo, the roasted the raw molybdenite concentrate. Molybdenite (MoS ) in the raw material disappeared, whereas products were leached at 60 °C for 3 h with 7 mol/L aqueous2 ammonium solution and liquid-solid molybdite (MoO ) was formed according to the following equations [26,27]. Meanwhile, rheniite ratio of 10 [4]. The3 results indicate that only 79.3% Mo was leached from the calcine of molybdenite (ReS ) was also oxidized to rhenium (VII) oxide (Re O ) whose vapor pressure reached 1 atm at concentrate2 without pretreatment. In contrast, the Mo 2leaching7 ratio reached 91.5% for the calcine of concentrate pretreated by acid leaching. This is attributed to the elimination of adverse effects of molybdates, which were insoluble in the aqueous ammonium solution. From the above results, it can be concluded that acid leaching as a method of molybdenite concentrate pretreatment not only played a significant role in improving the Re volatilization, but also had a crucial impact on the subsequent leaching of Mo.

4. Conclusions Separation of rhenium from a kind of lead-rich molybdenite concentrate by oxidative roasting was studied. The volatilization of Re and leaching ratio of Mo in ammonia solution were improved significantly after the lead and calcium components in the raw material were removed by hydrochloric acid. The results of acid leaching show that 93.6% of Pb and 97.4% of Ca were removed under the conditions of 95 °C for 10 min with HCl concentration of 8 wt. % and liquid-solid ratio of 5 (mL/g). The oxidative roasting experiment indicates the rhenium volatilization of 89.3% was obtained for the raw concentrate when roasted at 600 °C for 120 min in air. In contrast, the rhenium volatilization was enhanced distinctly to 98.0% when the pretreated concentrate was roasted at 550 °C for 100 min. Furthermore, the molybdenum leaching ratio of 91.5% was achieved by leaching the calcine of pretreated concentrate in aqueous ammonia solution. Conversely, only 79.3% of Mo was leached from the calcine of molybdenite concentrate without pretreatment. The XRD analysis confirms that the formation of molybdates was adverse for the rhenium separation and the subsequent leaching of molybdenum, which can be prevented after acid pretreatment.

Metals 2016, 6, 282 10 of 12

362.4 ◦C. Thus, rhenium oxide is absent in Figure 10, while molybdite was retained in the calcine during oxidative roasting: 2MoS2 + 7O2 → 2MoO3 + 4SO2 (6)

4ReS2 + 15O2 → 2Re2O7 + 8SO2 (7) By roasting the raw concentrate at 600 ◦C for 120 min under air atmosphere, galena (PbS) was oxidized to lead oxide (PbO), which combined with molybdite to form wulfenite (PbMoO4). In the meantime, calcite (CaCO3) in the raw material was decomposed to calcium oxide (CaO). Similarly, CaO would also react with molybdite to generate powellite (CaMoO4), as described by the following equations: 2PbS + 3O2 → 2PbO + 2SO2 (8)

PbO + MoO3 → PbMoO4 (9)

CaCO3 → CaO + CO2 (10)

CaO + MoO3 → CaMoO4 (11) As a result of hydrochloric acid leaching, the peaks of galena and calcite were disappeared. No obvious diffraction peaks of wulfenite and powellite except those of molybdite are observed in the XRD pattern of the roasted product, which was obtained at 550 ◦C for 100 min under air atmosphere. To evaluate the effect of removing lead and calcium on the leaching of Mo, the roasted products were leached at 60 ◦C for 3 h with 7 mol/L aqueous ammonium solution and liquid-solid ratio of 10 [4]. The results indicate that only 79.3% Mo was leached from the calcine of molybdenite concentrate without pretreatment. In contrast, the Mo leaching ratio reached 91.5% for the calcine of concentrate pretreated by acid leaching. This is attributed to the elimination of adverse effects of molybdates, which were insoluble in the aqueous ammonium solution. From the above results, it can be concluded that acid leaching as a method of molybdenite concentrate pretreatment not only played a significant role in improving the Re volatilization, but also had a crucial impact on the subsequent leaching of Mo.

4. Conclusions Separation of rhenium from a kind of lead-rich molybdenite concentrate by oxidative roasting was studied. The volatilization of Re and leaching ratio of Mo in ammonia solution were improved significantly after the lead and calcium components in the raw material were removed by hydrochloric acid. The results of acid leaching show that 93.6% of Pb and 97.4% of Ca were removed under the conditions of 95 ◦C for 10 min with HCl concentration of 8 wt. % and liquid-solid ratio of 5 (mL/g). The oxidative roasting experiment indicates the rhenium volatilization of 89.3% was obtained for the raw concentrate when roasted at 600 ◦C for 120 min in air. In contrast, the rhenium volatilization was enhanced distinctly to 98.0% when the pretreated concentrate was roasted at 550 ◦C for 100 min. Furthermore, the molybdenum leaching ratio of 91.5% was achieved by leaching the calcine of pretreated concentrate in aqueous ammonia solution. Conversely, only 79.3% of Mo was leached from the calcine of molybdenite concentrate without pretreatment. The XRD analysis confirms that the formation of molybdates was adverse for the rhenium separation and the subsequent leaching of molybdenum, which can be prevented after acid pretreatment.

Acknowledgments: This work was partially supported by the Co-Innovation Center for Clean and Efficient Utilization of Strategic Metal Resources, the Shenghua Lieying Program of Central South University under Grant 502035001, and the Innovation-Driven Program of Central South University under Grant 2016CXS021. Author Contributions: Guanghui Li and Zhiwei Peng conceived and designed and performed the experiments; Zhixiong You, Hu Sun and Rong Sun performed the experiments, analyzed the data and wrote the paper; Yuanbo Zhang and Tao Jiang supervised experimental work and reviewed the manuscript. Conflicts of Interest: The authors declare no conflict of interest. Metals 2016, 6, 282 11 of 12

References

1. Hu, K.; Zhao, D.; Liu, J. Synthesis of nano-MoS2/bentonite composite and its application for removal of organic dye. Trans. Nonferr. Met. Soc. 2012, 22, 2484–2490. [CrossRef] 2. U.S. Geological Survey. Minerals Commodity Summaries (2015) and Minerals Yearbook (2012). Available online: http://minerals.usgs.gov/minerals/pubs/commodity/rhenium/mcs-2012-rheni.pdf (accessed on 23 March 2015). 3. Mccandless, T.; Ruiz, J.; Campbell, A. Rhenium behavior in molybdenite in hypogene and near-surface environments: Implications for Re-Os geochronometry. Geochim. Cosmochim. Acta 1993, 57, 889–905. [CrossRef] 4. Xiang, T. Molybdenum Metallurgy; Central South University Press: Changsha, China, 2002; pp. 23–34. (In Chinese) 5. Tarasov, A.; Besser, A.; Gedgagov, E. Integrated technology for processing rhenium-containing molybdenite concentrates to recover molybdenum and rhenium into commercial products. Miner. Process. Extr. Metall. Rev. 2001, 22, 509–517. [CrossRef] 6. Shariat, M.; Hassani, M. Rhenium recovery from Sarcheshmeh molybdenite concentrate. J. Mater. Process. Technol. 1998, 74, 243–250. [CrossRef] 7. Tripathy, P.; Rakhasia, R. Chemical processing of a low grade molybdenite concentrate to recover molybdenum. Trans. Inst. Min. Metall. Sect. C 2006, 115, 8–14. [CrossRef] 8. Juneja, J.; Singh, S.; Bose, D. Investigations on the extraction of molybdenum and rhenium values from low grade molybdenite concentrate. Hydrometallurgy 1996, 41, 201–209. [CrossRef] 9. Ketcham, V.; Hazen, W.; Coltrinari, E. Pressure oxidation process for the production of from molybdenite. U.S. Patent US614988321, November 2000. 10. Khoshnevisan, A.; Yoozbashizadeh, H.; Mozammel, M.; Sadrnezhaad, S. Kinetics of pressure oxidative leaching of molybdenite concentrate by . Hydrometallurgy 2012, 111–112, 52–57. [CrossRef] 11. Jiang, K.; Wang, Y.; Zou, X.; Zhang, L.; Liu, S. Extraction of molybdenum from molybdenite concentrates with hydrometallurgical processing. JOM 2012, 64, 1285–1289. [CrossRef] 12. Joo, S.; Kim, Y.; Kang, J.; Yoon, H.; Kim, D.; Shin, S. Recovery of molybdenum and rhenium using selective precipitation method from molybdenite roasting dust in alkali leaching solution. Mater. Trans. 2012, 53, 2038–2042. [CrossRef] 13. Cao, Z.; Zhong, H.; Liu, G.; Fu, J.; Wang, S.; Qiu, Y. Electric-oxidation kinetics of molybdenite concentrate in acidic NaCl solution. Can. J. Chem. Eng. 2009, 87, 939–944. 14. Molaei, R.; Shariat, M.; Zakeri, A.; Boutorabi, M. Rhenium (VII) recovery via electro-oxidation process I. The influence of time, temperature, and pH. Electrochem. Solid. St. Lett. 2011, 14, E1–E3. [CrossRef] 15. Askari Zamani, M.; Hiroyoshi, N.; Tsunekawa, M.; Vaghar, R.; Oliazadeh, M. Bioleaching of Sarcheshmeh molybdenite concentrate for extraction of rhenium. Hydrometallurgy 2005, 80, 23–31. [CrossRef] 16. Olson, G.; Clark, T. Bioleaching of molybdenite. Hydrometallurgy 2008, 93, 10–15. [CrossRef] 17. Peng, Z.; Hwang, J. Microwave-assisted metallurgy. Int. Mater. Rev. 2015, 60, 30–63. [CrossRef] 18. Joo, S.; Kim, Y.; Kang, J.; Kumar, J.; Yoon, H.; Parhi, P. Recovery of rhenium and molybdenum from molybdenite roasting dust leaching solution by ion exchange resins. Mater. Trans. 2012, 53, 2034–2037. [CrossRef] 19. Lan, X.; Liang, S.; Song, Y. Recovery of rhenium from molybdenite calcine by a resin-in-pulp process. Hydrometallurgy 2006, 82, 133–136. [CrossRef] 20. Seo, S.; Choi, W.; Yang, T.; Kim, M.; Tran, T. Recovery of rhenium and molybdenum from a roaster fume scrubbing liquor by adsorption using activated . Hydrometallurgy 2015, 129–130, 145–150. [CrossRef] 21. Keshavarz Alamdari, E.; Darvishi, D.; Haghshenas, D.; Yousefi, N.; Sadrnezhaad, S. Separation of Re and Mo from roasting-dust leach-liquor using solvent extraction technique by TBP. Sep. Purif. Technol. 2012, 86, 143–148. [CrossRef] 22. Marin, T.; Utigard, T.; Hernandez, C. Roasting kinetics of molybdenite concentrates. Can. Metall. Q. 2009, 48, 73–80. [CrossRef] 23. Mchugh, L.; Balliett, R.; Mozolic, J. The sulfide ore looping oxidation process: An alternative to current roasting and smelting practice. JOM 2008, 60, 84–87. [CrossRef] Metals 2016, 6, 282 12 of 12

24. Kim, B.; Lee, H.; Choi, Y.; Kim, S. Kinetics of the oxidative roasting of low grade Mongolian molybdenite concentrate. Mater. Trans. 2009, 50, 2669–2674. [CrossRef] 25. Wang, L.; Zhang, G.; Dang, J.; Chou, K. Oxidation roasting of molybdenite concentrate. Trans. Nonferr. Met. Soc. 2015, 25, 4167–4174. [CrossRef] 26. Shigegaki, Y.; Basu, S.; Wakihara, M.; Taniguchi, M. Thermal analysis and kinetics of oxidation of molybdenum sulfides. J. Therm. Anal. Calorim. 1988, 34, 1427–1440. [CrossRef] 27. Utigard, T. Oxidation mechanism of molybdenite concentrate. Metall. Mater. Trans. B 2009, 40, 490–496. [CrossRef] 28. Ma, G.; Lei, N.; Wang, Z.; Wang, W.; Feng, B. Research on reducing the lead content of high-lead molybdenite concentrates. China Molybd. Ind. 2015, 39, 23–25. (In Chinese) 29. Gupta, C. Extractive Metallurgy of Molybdenum; CRC Press: Boca Raton, FL, USA, 1992; pp. 156–157. 30. Holdich, R.; Lawson, G. The solubility of aqueous lead chloride solutions. Hydrometallurgy 1987, 19, 199–208. [CrossRef]

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