extractive metallurgy of rhenium

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MINERALS & METALLURGICAL PROCESSING Vol. 30 No. 1 February 201359

Extractive metallurgy of rhenium: a reviewC.D. Anderson, P.R. Taylor and C.G. AndersonPhD student, professor and professor, respectively, Kroll Institute for Extractive MetallurgyColorado School of Mines, Golden, CO

AbstractA variety of processing technologies exist for recovery from both primary and secondary sources of rhenium. Currently, there are no known primary rhenium deposits; thus, the method in which primary rhenium is produced is dependent on the commodity of which it is a byproduct, e.g., copper, molybdenum, uranium, etc. In addition, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing. This paper presents a review of both primary and secondary processing technologies for the recovery of rhenium.

Paper number MMP-12-077. Original manuscript submitted November 2012. Revised manuscript accepted for publication December 2012. Discussion of this peer-reviewed and approved paper is invited and must be submitted to SME Publications Dept. prior to August 31, 2013. Copyright 2013, Society for Mining, Metallurgy, and Exploration, Inc.

IntroductionIn recent times, the aerospace and petrochemical

industries have come to rely on a silver-white transition metal, with a specific gravity of 21 and the second-highest melting point (3,180 C) of any metal in the periodic table. This metal is rhenium (Re), the last natural element to be discovered, in 1925, by Mr. and Mrs. Walter Noddack and Professor Otto Berg. The element was named after the Rhine River of their na-tive Germany (Millensifer, 2010). Rheniums unique properties have made it a vital part of the superalloy industry, most prominently in nickel superalloys used in both aerospace and industrial gas-fired turbines. Its second largest application is in the catalyst industry for the production of unleaded gasoline (Polyak, 2011).

The concentration of rhenium in the earths crust is rather small, at approximately 0.7-1.0 parts per billion (Millensifer, 2010), although there is specu-lation that this number may be as high as 10 ppb (Fleischer, 1959). The only documented occurrence of rhenium as a mineral, rheniite (ReS2), was found near the Russian Kudryavyi volcano (Korzhinsky et al., 1994; Tessalina et al., 2008). Other than this specific occurrence, rhenium is typically associated with molybdenum-copper porphyry deposits in con-centrations of up to 0.2% (Woolf, 1961). Currently, there are no producers of primary rhenium, and almost all rhenium, with few exceptions, is produced as a byproduct of the molybdenum and copper industries.

In 2011, Chile produced the most rhenium (~25,000 kg) and contained the worlds largest rhenium reserve (~1.3 M kg) (Polyak, 2011). One processor in particu-

lar, Chiles Molibdenos y Metales S.A. (Molymet), is the largest molybdenum and rhenium producer in the world (Whittaker, 2012). The 2011 geographic breakdown of worldwide rhenium reserves and production quantities are shown in Table 1.

Rhenium is typically sold on long-term contracts between consumers and producers. Unfortunately, data available on historical rhenium prices is mostly based on free market pric-ing and, thus, may not reflect the actual pricing of the rhenium market to the desired extent. Nevertheless, Fig. 1 illustrates a combination of historical free market rhenium prices from a variety of sources, which demonstrate price fluctuations in the market. Some of the major fluctuations may be attributed to the following historical events (Polyak, 2011; Naumov, 2007; Blossom, 1998):

1970: Use of rhenium in petrochemical catalysts begins. 1980: Amount of rhenium used in petrochemical cata-

lysts doubles. 1991: Dissolution of U.S.S.R. leads to an increased sup-

ply of rhenium in the market. 1999, 2003, 2006: Appearance of new generation turbine

blades containing rhenium.

As of 2011, the U.S. Geological Survey (USGS) reported that the free market price for 99.99% pure rhenium metal powder is approximately $2,000/kg (Polyak, 2011).

Due to the lack of primary rhenium deposits, the method with which it is processed is directly related to the method with which the minerals it is associated with are produced, typically copper and molybdenum. Characteristically, rhenium found in mixed copper/molybdenum deposits is first separated with the molybdenum from the copper, using conventional

Key words: Rare-earth minerals, Rhenium, Ammonium perrhenate, Extractive metallurgy

Minerals & Metallurgical Processing, 2013, Vol. 30, No. 1, pp. 59-73. An official publication of the Society for Mining, Metallurgy, and Exploration, Inc.

SPECIAL RARE-EARTH MINERALS ISSUE

February 2013 Vol. 30 No. 1 MINERALS & METALLURGICAL PROCESSING60

concentration technologies, typically froth flotation. After this, rhenium is recovered as a byproduct from molybdenum processing (Naumov, 2007).

The principal molybdenum end product, molybdenum tri-oxide (MoO3), is the basic raw material for most commercially used products of molybdenum. Therefore, the method in which rhenium is produced is dependent on the production method of MoO3. Molybdenum trioxide is produced via pyrometallurgi-cal roasting or hydrometallurgical pressure oxidation (Gupta, 1992; Ketcham et al., 2000).

In addition to being produced as a byproduct of the mo-lybdenum/copper industry, focus on the recovery of rhenium from secondary sources, such as alloy scraps and catalysts, is continually growing.

Accordingly, the following sections are broken into both the methods by which rhenium is recovered as a byproduct of primary processing, as well as from secondary sources.

Rhenium recovery from pyrometallurgical effluent streams

In pyrometallurgical roasting of molybdenum concentrates containing rhenium, rhenium present in the molybdenum concentrate is oxidized to rhenium heptoxide (Re2O7) via the following simplified reaction.

2ReS2 +7.5O2(g) = Re2O7(g)+ 4SO2(g)

(G298K = - 406.5 kcal) (1)

It should be noted that rhenium heptmoxide is extremely volatile (Pvap = 711 mmHg at 633K); thus, at the temperatures used for molybdenum roasting (900-950 K) it is likely that nearly all of the rhenium present is volatilized. This volatile product exits the furnace with the flue gases and is subsequently recovered as perrhenic acid (HReO4) after being scrubbed with water via the following reaction.

Re2O7(g) + H2O = 2HReO4(aq)

(G298K = -15.16 kcal) (2)

After scrubbing, the aqueous rhenium is typically recovered using solvent extraction or ion-exchange processes. Gener-ally, the end precursor product produced by these methods is ammonium perrhenate (NH4ReO4) via crystallization (Mil-lensifer, 2010).

In the following subsections, rhenium recovery methods that rely on the pyrometallurgical volatilization of rhenium and production of an ammonium perrhenate as an intermediate product are discussed and broken into their respective branches of concentration technology.

Ion exchange processes. The original Kennecott Process for the recovery of rhenium involves the previously mentioned scrubbing process, in which the exhaust gases of the molyb-denum roasting circuit are washed with water to produce per-rhenic acid. The pregnant solution is continually recirculated through the scrubber circuit until the rhenium concentration reaches approximately 100 mg/L. After this, the solution is conditioned for 24 hours with caustic soda, soda ash and oxidized with calcium hypochlorite. The pH of this solution is brought up to 10 to precipitate any contaminants (primar-ily iron) remaining in solution and allow them to settle. The solution is filtered and sent to an ion exchange circuit, where the anionic resin preferentially adsorbs the aqueous rhenium from the alkaline solution. After the loading stage is finished, the rhenium is stripped by the addition of hydrochloric acid. A caustic soda solution is then used to remove the remaining adsorbed molybdenum, which is subsequently recovered as calcium molybdate. Perchloric acid and hydrogen sulfide are then added to this solution to precipitate rhenium as rhenium

Figure 1 Historical rhenium metal power price from 1952-2012 (year 2012: Metal Pages 2012; years 2007-2011: Polyak, 2011; years 1999-2006: Naumov, 2007; years 1952-1998: Blossom, 1998).

Table 1 Worldwide rhenium reserves and production by country. Source: Polyak, 2011.

Country Reserves (kg) Production (kg)

Chile 1,300,000 25,000

United States 390,000 6,000

Russia 310,000 1,500

Kazakhstan 190,000 2,500

Armenia 95,000 400

Peru 45,000 5,000

Canada 32,000 1,800

Poland NA 4,500

Other countries 91,000 1,500

World total (rounded) 48,000 2,500,000

NA = not available


sulfide (Re2S7). This precipitate is then redissolved in a solution of ammonia and hydrogen peroxide, from which it crystallizes as ammonium perrhenate (NH4ReO4). Figure 2 illustrates the flowsheet for the original Kennecott Process.

Currently, KGHMs Glogow smelting facility (Glogow, Po-land) is utilizing ion exchange technology (the Ecoren process) for the recovery of aqueous rhenium, from copper concentrate smelter flue gas. In this process, the rhenium within the copper concentrate is volatilized as rhenium heptoxide and reports to the sulfuric acid scrubbing plant. Like the Kennecott Process, the rhenium heptoxide is scrubbed using water, and brought into solution along with various other impurities, including molybdate, sulfate and sulfite salts. The typical rhenium con-centration in this solution is 0.02 g/L. This solution is sent to a closed polypropylene filter press to remove any residual silicon prior to being sent to the ion exchange columns. The

solution is then sent to a series of ion exchange columns, where rhenium is selectively adsorbed on the weakly basic anionic resin as the perrhenate anion (ReO4

-). After the loading phase, the rhenium is eluted using an ammonia solution. The loaded eluent is the sent to vacuum crystallization, where ammonium perrhenate is produced. Typically, repeated recrystallization is necessary to produce a high-grade (99.95% pure) ammonium perrhenate product. The Glogow smelter facility has the ca-pability to produce 4-5 t of ammonium perrhenate per year. The Ecoren process flowsheet is shown in Fig. 3 (Chmielarz and Litwinionek, 2010).

Solvent extraction processes. Some of the richest rhenium-containing deposits in the world are found in Kazakhstan. In the Zhezkhazgan deposit, copper concentrates produced can contain up to 30 g/t of rhenium. Abisheva et al. have proposed

Figure 2 The original Kennecott Process (Sutulov, 1965).


a process in which the copper concentrate is electrosmelted, and the rhenium present is volatilized, as rhenium heptoxide (Re2O7), and scrubbed with water/dilute sulfuric acid to form perrhenic acid (HReO4) in concentrations of up to 0.25 g/L. The aqueous rhenium solution is sent to a solvent extraction step for the selective separation of rhenium from the impurity elements present. The extractant used in this technology is tri-alklamine organic compound (TAA) in a kerosene diluent. The loaded organic phase is stripped using ammonium hydroxide to produce ammonium perrhenate. The barren organic phase is then recycled for further use in solvent extraction, and the impure ammonium perrhenate is continually dissolved and recrystallized to produce a 98.5% pure ammonium perrhenate product (Abisheva et al., 2011). A flowsheet of this process is shown in Fig. 4.

The use of a tertiary amine rhenium extractant was con-

firmed by Singh et al., who ran a 250-L rhenium extraction pilot plant using it. Results showed that 98% of the aqueous rhenium was recoverable, and allowed for the production of a 98% pure rhenium precursor product using solvent extraction (Singh et al., 1982).

As an alternative to the conventional method of ammonium perrhenate precipitation, the USBM developed a solvent-extraction/electrowinning process for rhenium recovery from molybdenite roasting. Once more, rhenium heptoxide present in the flue gas was scrubbed using water. The aqueous rhenium was oxidized using sodium chlorate and the pH of the solu-tion was brought to 12 by caustic soda addition. Prior to the solvent extraction step, the solution was filtered to remove any precipitates that formed during pH adjustment. The pregnant solution was then sent to a six-stage extraction circuit, which used a combination of: 5% aliquat 336, 5% primary decyl

Figure 3 KGHM Ecoren rhenium recovery at the Glogow smelter (Chmielarz and Litwinionek, 2010).


alcohol (PDA) and 90% kerosene as the extractant phase. The loaded organic was then stripped using a 1-M perchloric acid/ammonium sulfate solution. After stripping, the rhenium-rich electrolyte was sent for electrowinning using a current density of 360 A/m2. Results from the pilot plant experimentation show that rhenium metal can be prepared from dilute impure so-lutions containing aqueous rhenium and molybdenum (Churchward and Rosenbaum, 1963). An illustration of the flowsheet for this process is shown in Fig. 5.

Alternative methods for rhenium recovery In addition to the use of conventional ion exchange

or solvent extraction processes, there are a number of processes that involve novel methods for the extraction and production of rhenium. The following section will describe each of these in their respective context.

Recovery from Mo/Cu ores. As an alternative to traditional roasting technology, Rio Tintos Ken-necott facility at Bingham Canyon, UT has patented (US# 6149883) an alkaline pressure oxidation process for the recovery of molybdenum and rhenium from molybdenite concentrates (molybdenum autoclave process or MAP). In this process, the molybdenite flotation concentrate is leached with either sodium or potassium hydroxide at elevated temperature (150-200 C) and pressure (517 - 1,400 kPa or 75-200 psig) to form soluble molybdate (MoO4

2-).The aqueous molybdate is then recovered using solvent extraction and is stripped with an ammonium hydroxide eluent. The rhenium present in the molybdenite concentrate is primarily recovered in the solvent extraction step, in the loaded strip solution. After this, it is recovered using a selective ion exchange resin, which typically contains quaternary amine functional groups (Ketcham et al., 2000).

In 2009, Freeport-McMoRan applied for a patent Figure 5 USBM SX/EW process flowsheet for rhenium recovery (Churchward and Rosenbaum, 1963).

Figure 4 Flowsheet for the recovery of rhenium from sulfuric acid scrubbing solution (Abisheva et al., 2011).


(US# 0263490A1) for a proprietary method of recovering rhenium as a byproduct of copper leaching/molybdenum oxi-dation. The application states that, the rhenium rich PLS can originate from an active copper leach, stockpile copper leach, acid blowdown stream, or a leach of molybdenite roaster flue fumes and dusts. This loaded stream is sent to an activated carbon column circuit for adsorption of the aqueous rhenium. After adsorption, the loaded rhenium is stripped using an elu-ent solution containing approximately 2.5% sodium hydroxide and 2.5% ammonium hydroxide at a pH of 7, and an operat-ing temperature of 80-110 C. After elution, the rhenium-rich stream is sent for rhenium recovery to produce a pure rhenium product, and a rhenium lean eluate solution for optional reuse in the circuit (Waterman et al., 2009). An example of Freeports proposed rhenium recovery circuit is shown in Fig. 6.

In October 2012, the Australian Patent Office granted a

patent (AUS# 2011229125) to Alexander Mining Plc. for the MoReLeach process (Sutcliffe et al., 2012). Officially titled Method of oxidative leaching of molybdenum-rhenium sul-fide ores and/or concentrates, the process involves leaching a molybdenum/rhenium concentrate in a closed reactor vessel at atmospheric temperature and pressure. The lixiviant used is an aqueous solution of chlorine-based oxidizing species, in which the predominant chlorine-based oxidizing species are hypochlorite (ClO-) ions. The molybdenum sulfide is oxidized to the soluble molybdate anion (MoO4

2-), and the rhenium present is oxidized to the perrhenate anion (ReO4

-) via the following proposed reactions (Gupta, 1992). Note: the second reaction is proposed by the author.

MoS2 + 9NaClO + 3H2O = MoO42-

(aq) + 9NaCl +

2SO42-

(aq) + 6H+ (G298K = -506.9 kcal) (3)

Figure 6 Freeport pressure oxidation rhenium recovery process (Waterman et al., 2009).


sorbant, activated charcoal in the gold industry. But, in this process, an ion exchange resin is added to a solution that has already undergone leaching. Thus, after perrhenic acid is produced by scrubbing with water, the ion exchange resin is added to selectively remove the rhenium and molybdenum from solution, via adsorption onto the resin. Experimental results show that by using a strong basic anionic exchange resin, rhenium can be selectively adsorbed from solution. Ad-ditionally, elution results have shown that molybdenum can be selectively eluted from the resin by using ammonium chloride (NH4Cl) as the eluent, leaving rhenium adsorbed. After this, rhenium can then be stripped from the resin using dilute nitric acid (Lan et al., 2006).

In 1947, Melaven and Bacon, of the University of Tennes-see, patented (US# 2,414,965) a technology for the recovery of rhenium from molybdenite roasting flue dust. In their process, cyclone flue dust was recovered, and rhenium was leached with water and compressed air. After lixiviation, potassium chloride was added to the solution to recover rhenium as a pure potassium perrhenate (KReO4) precipitate. Metallic rhenium was then produced by roasting the potassium perrhenate in a silver tube under a hydrogen atmosphere at 350 C. After roasting, the residue is washed with hot distilled water, leav-ing a pure metallic rhenium product (Melaven and Bacon,

Figure 7 MoRe leach process (Sutcliffe et al., 2012).

ReS2+ 9.5NaClO + 2.5H2O = ReO4-(aq) + 9.5NaCl +

2SO42-

(aq) + 5H+ (G298K = -535.7 kcal) (4)

After dissolution, the pregnant solution is separated from the undissolved residue and sent for a metal separation stage such as solvent extraction, ion exchange, etc. Additionally, the proposed process allows for the regeneration of hypochlorite for subsequent reuse as a reagent. A flowsheet of the proposed process is illustrated in Fig. 7.

In the mid-1990s, after discovering rheniite (ReS2) at the Russian Kudryavyi volcano, Russian researchers found that the high-temperature fumarole gases exiting the volcano contained considerable (0.5-2.5 g/t) of rhenium in the form of gaseous rhenium chlorides (ReCl5) and fluorides (ReF5). Currently, researchers are working on using zeolites as the adsorption medium for the rhenium rich fumarole gases. As of 2007, re-searchers are operating this process on a pilot scale, although this process may prove rather difficult, due to the inherent danger of the natural environment and lack of infrastructure within the area (Naumov, 2007; Sinegribov et al., 2007).

Lan et al. have proposed a novel process for the recovery of rhenium-containing flue dusts, using an application commonly used in gold hydrometallurgy, known as the resin-in-pulp process. The resin-in-pulp process relies on the addition of a


1947; Sutulov, 1965). The flowsheet for the Melaven process is shown in Fig. 8.

Recovery from uranium leach liquors. After finding rhenium adsorbed on their uranium ion-exchange columns in Palangana, TX, efforts were made for the recovery of rhenium as rhenium sulfide. Rhenium was present in the ammonium carbonate leach liquor as the perrhenate anion, ReO4

-, which is selectively adsorbed on the anionic exchange resin. The proposed process calls for the elution of this anion by way of ammonium nitrate and precipitation of rhenium sulfide (Re2S7) with the addition of hydrogen sulfide gas (Goddard, 1996).

The proposed flowsheet for this process is shown in Fig. 9. In another case where rhenium is found in uranium leach-

ing solutions, Chekmarev et al. have proposed the recovery of rhenium using complexation and ultrafiltration methods with the aid of water-soluble polyelectrolytes. Rhenium in the pregnant leach solution is complexed using VA-type cationic polyelectrolytes containing quaternary ammonium base groups. These complexes are sent through ultrafiltration to selectively remove the high molecular weight rhenium complex, leaving a wash solution ready for subsequent rhenium processing (Chekmarev et al., 2004).

Figure 8 The Melaven process (Melaven and Bacon, 1947).


Recycling of rheniumIn addition to being produced as a byproduct of molybde-

num, it is possible to recycle rhenium during processing, and after its use within the industry. The following subsections will illustrate a number of the current and proposed technologies that might be used to recover rhenium from secondary sources.

Recycling from superalloys and alloy scraps. Due to the

Figure 10 Schematic of the W-Re tube furnace (Heshmatpour and McDonald, 1982).

high temperature properties of tungsten and rhenium, it is not uncommon to find these metals in alloys together. Thus, W-Re scrap may be recycled via an oxidative pyrometallurgical roasting technique. Initially, the scrap is roasted at 1,000 C, under an oxidizing atmosphere to produce rhenium heptoxide (Re2O7), which is subsequently condensed in the cooler part of the tube furnace (Fig. 10).

This material is then sent for digestion in water. The aque-

Figure 9 Recovery of rhenium sulfide (Goddard, 1996).


ous rhenium (ReO4-) is subsequently precipitated as potassium

perrhenate upon the addition of potassium chloride via the following reaction.

KCl + ReO4- = KReO4 + Cl

-(aq)

(G298K = -6.133 kcal) (5)

The potassium perrhenate is filtered and further purified via continued dissolution and recrystallization. After purifica-tion, the salt is dried and sent for reduction under a hydrogen atmosphere at approximately 350 C. Experimental results show that 93.1% of the rhenium was recovered to produce a 99.98% pure Re product (Heshmatpour and McDonald, 1982).

H.C. Starck has applied for a patent (US # 0255372 A1) for a process for the elevated temperature digestion and recycling of rhenium-containing superalloys. Initially, the superalloys are digested in a molten salt melt containing NaOH, Na2CO3, and Na2SO4 at temperatures of 850-1,100 C in a directly fired rotary kiln. In addition to this, oxidizing agents such as nitrates and peroxides of the alkali metals are added. The melt from this process is then cooled and sent to a comminution process for size reduction. The material is then leached using water as the lixiviant to dissolve the 6 and 7th group elements present in the superalloy. This slurry is then filtered to separate the insoluble

Figure 11 The H.C. Starck process for superalloy recycling (Olbrich et al., 2009).

Co, Ni, Fe, Mn and Cr from the leach liquor. Magnetic sepa-ration is then applied to the insoluble components for further separation and concentration. The pregnant leach solution is sent to an ion exchange step, where the aqueous rhenium is selectively adsorbed and can be recovered using the methods described in the previous sections (Olbrich et al., 2009). An example of the flowsheet for this process is shown in Fig. 11.

Through the use of electrolytic decomposition of rhenium superalloys, Stoller et al. have patented (US# 0110767) a process that involves the use of titanium baskets as electrodes. The baskets containing the superalloy scrap are fed to a poly-propylene electrolysis cell containing a 18% HCl solution. The electrolytic dissolution is carried out for 25 hours at a frequency of 0.5 Hz, current of 50 amps, voltage of 3-4 V and a temperature of 70 C. The remaining scrap is then filtered from the pregnant solution and sent for further dissolution in sodium hydroxide/peroxide solution. After completion, this filtrate is sent to ion exchange for the recovery of rhenium and molybdenum. Rhenium is recovered using the ion exchange processes discussed previously (Stoller et al., 2008). An il-lustration of this process is shown in Fig. 12.

Recycling of spent Pt-Re catalystsPetroleum-reforming catalysts containing rhenium and plati-


Figure 12 Electrochemical method for recycling of superalloys (Stoller et al., 2008).

num on an alumina substrate are used in the refining industry for the improvement of the octane level of fuels. After being deactivated, an effective method for the recovery or rhenium and other PGM metals is necessary. There are two basic meth-ods by which this is achieved (Kasikov and Petrova, 2009):

1. Complete dissolution of the alumina substrate.2. Selective dissolution and recovery of rhenium and

platinum.

The following summaries will present examples of both of these technologies; for further information, refer to Kasikovs literature review Processing of deactivated platinum-rhenium catalysts (Kasikov and Petrova, 2009).

Complete dissolution of the alumina substrate. During complete dissolution of the alumina substrate, sulfuric acid may be used for dissolution of alumina, rhenium and, to some extent, platinum. The rhenium-rich solution is separated from the platinum-containing residue and aqueous aluminum us-ing ion exchange. Rhenium is eluted from the organic amine resin by way of hydrochloric acid addition. After elution, the rhenium-rich eluate is neutralized using ammonium hydrox-ide. This solution is then evaporated to form a super-saturated

solution, and cooled to allow for crystallization of ammonium perrhenate. After continued redissolution and recrystallization, a high-purity ammonium perrhenate precipitate is produced (El Guindy, 1997). The flowsheet for this process is shown in Fig. 13.

As an alternative to sulfuric acid, sodium bicarbonate may also be used as a lixiviant. The proposed advantage of this process is the complete removal of the ion exchange circuit shown in Fig. 14. Experiments were performed on crushed and uncrushed catalysts in both packed columns and agitated leach vessels. Experimental results showed that that rhenium is preferentially leached in the sodium bicarbonate solution. Rhenium recovery reached 97% for crushed catalysts, and 87% for uncrushed catalyst samples. After dissolution, the aque-ous rhenium is crystallized via evaporative crystallization as an ammonium perrhenate intermediate product (Angelidis et al., 1999). The flowsheet for this process is shown in Fig. 14.

Selective leaching of rhenium and platinum. The methods used to selectively recover platinum and rhenium from spent catalysts without completely dissolving the alumina substrate vary from calcination of the catalysts to selective leaching in alkaline or acid conditions at ambient and elevated temperatures (Kasikov and Petrova, 2009).


Figure 13 Method for rhenium recovery from spent reforming catalysts (El Guindy, 1997).


By calcining the catalyst at temperatures up to 1,150 C, the -Al2O3 undergoes a phase transition to the chemically stable -Al2O3 phase, lowering the dissolution of the alumina catalyst. The platinum and rhenium can then be selectively leached in concentrated (5 mol/L) sulfuric acid solutions containing sodium chloride and a potassium persulfate oxidant (K2S2O8). Kpumaneva reports that rhenium and platinum recoveries are as high as 95.5% and 97%, respectively (Kpumaneva et al., 2001).

Additionally, a U.S. patent (US#: 5542957) has been granted involving the selective leaching of platinum and rhenium at elevated temperatures (50-300 C) and pressures (207-9,000 kPa or 30-1,300 psig). In this process, a dilute solution of sulfuric acid (0.001-1.0 mol/L) is used in the presence of ammonium

iodide or bromide and oxygen to selectively leach Pt and Re, while leaving behind the alumina substrate. At 160 C and an oxygen overpressure of 800 kPa (116 psig), the authors report a rhenium recovery of 98% (Han and Meng, 1996).

Production of metallic rheniumGenerally, metallic rhenium is not produced at the facility

at which it is concentrated and separated from other elements. Instead, it is produced from ammonium perrhenate (APR) using methods similar to those used in the molybdenum and tungsten industries; i.e., a reductant such as carbon monoxide or hydrogen is used. Although APR is the typical precursor material that is reduced, potassium perrhenate may also be

Figure 14 - Rhenium recovery using sodium bicarbonate (Angelidis et al., 1999).


used (Sutulov, 1965). The two primary methods for the production of metallic

rhenium are:

1. The reduction of ammonium perrhenate (NH4ReO4).2. The reduction of potassium perrhenate (KReO4).

Ammonium perrhenate (NH4ReO4) is the typical precursor product used in the production of metallic rhenium and rhenium compounds, including metallic rhenium powder and perrhenic acid. Metallic rhenium is produced by being reduced by hy-drogen gas at elevated temperature, T = 1,000 C (Hurd and Brimm, 1939). Ammonium perrhenate (APR) is placed in boats and subjected to countercurrent hydrogen gas flow. Depending on the particle size of metallic rhenium powder product, the reduction may be completed in single or multiple stages and the APR may be ground prior to reduction (Millensifer, 2010). The proposed reaction for this process is shown below. Figure 15 illustrates the typical production steps in making metallic rhenium from ammonium perrhenate.

2NH4ReO4(s) + 7H2 = 2Re + 2NH3 + 8H2O

(G298K = -37.16 kcal) (6)

In Hurd and Brimms process for the commercial reduction of potassium perrhenate, the feed material is crushed to ap-proximately 60 US mesh and dried at 175 C. The material is then placed in a silver boat inside of a refractory tube furnace. The boat is heated to 250 C under a hydrogen atmosphere for

Figure 15 Production of rhenium from ammonium perrhenate (Millensifer, 2010).

two hours, and then the temperature is raised to 500 C for an additional two hours. The reduction product is then cooled slowly in an inert atmosphere to prevent oxidation and washed with water to remove any residual alkaline material (Hurd and Brimm, 1939). The proposed reaction for this technique is shown below.

2KReO4 + 7H2 = 2Re + 2KOH + 6H2O (G298K = -16.07 kcal) (7)

ConclusionMost of the processes involved in the production of primary

and secondary rhenium involve the use of either elevated temperatures, elevated pressures, large amounts of reagents or a combination of the three. Thus, it is imperative that the extraction and recovery of rhenium is as efficient as economi-cally possible. Some possible research opportunities for the development of higher efficiency processes are suggested in the following paragraphs.

The roasting of molybdenum concentrates containing rhenium is an elevated-temperature, exothermic process that requires temperatures of 900-950 K. Early scrubbing attempts of flue gases were relatively inefficient, capturing roughly 25% of the rhenium present. Through innovation in the scrubbing equipment used, recoveries have now increased to approxi-mately 80% (Millensifer, 2010). Thus, increasing the recovery of rhenium from roasting flue dusts presents itself as a viable research opportunity.


Conversely, another option for increased rhenium recovery may involve the use of hydrometallurgical pressure oxidation of molybdenum concentrates. Although this process involves the use of elevated temperatures/pressures and additional re-agents, the enhanced recoveries inherent to the process may make this a viable alternative to traditional roasting processes. Research of this technique, and other technologies involving the use of low temperature hydrometallurgical oxidation is currently ongoing in industry.

As a complementary process to both roasting and pressure oxidation, the need for efficient separation of aqueous rhenium from the process stream is essential. A variety of techniques have been the subject of investigation, including ion exchange, solvent extraction and activated carbon. With the continual development of higher selectivity resins and extractants used in both ion exchange and solvent extraction, the area presents itself worthy of further research and development.

Relatively new to the rhenium industry is the processing of rhenium-laden manufacturing scrap and end-of-life materials, such as catalysts and super alloys. Since this is such a new faction of the rhenium industry, the research possibilities are almost endless, but should definitely include the utilization of primary processing techniques, as well as the implementation of end-of-life recycling programs.

As rhenium sources are depleted, the need for efficient and economical extraction from both primary and secondary sources is essential in maintaining the rhenium supply. This paper has presented a review of both the current and former technolo-gies used in the extractive metallurgy of rhenium, as well as provided some areas for potential technological development.

AcknowledgmentsSpecial recognition is due to Tom Millensifer for his assis-

tance with this paper; in addition, the authors thank the Office of Naval Research for the financial support that allowed this research to occur.

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