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By R. G. Benson, P. R. Wotruba, K. W. Schmidt

Introduction The Fortitude deposit is located in Nevada's Battle Mountain mining district in the central Great Basin region of the Basin and Range province of the western US. The deposit is at Copper Canyon in the southernmost portion of the district, about 19 km (12 miles) southwest of Battle Mountain (Fig. 1). Mining activity in the district began in 1866 (Roberts and Arnold, 1965). Base and precious metals were mined by underground methods intermittently until 1955. Placer gold was first discovered at Copper Canyon in 1912. Moderate scale dredging of alluvial gravel at the mouth of Copper Canyon was conducted by the Natomas Co. from 1944 to 1955 (Johnson, 1973). In 1967, Duval Corp. began large scale open-pit copper mining. (Sayers, Tippett, and Fields, 1968). Copper prices decreased and precious metals prices increased in the mid-1970s so Duval's exploration efforts at Copper Canyon shifted to precious metals. These exploration efforts resulted in the discovery of the Tomboy-Minnie deposits and shortly afterward, the Northeast Extension deposit at Copper Canyon (Blake, Wotruba, and Theodore, 1984; Fig. 2). The Fortitude deposit was discovered in late 1980, after Duval had proceeded with precious metals production at Copper Canyon. Converting the existing plant facilities at Copper Canyon to a precious metals recovery system was completed in 1978 (Jackson, 1982). Production from the Tomboy-Minnie deposits began in early 1979. Precious metals production moved to the Northeast Extension deposit in 1982. In late 1984, after three years of preproduction stripping and mining of the Upper Fortitude ore zone, production from the higher grade Lower Fortitude ore zone began. Currently, Battle Mountain Gold Co. operates the Fortitude deposit. Production from the Battle Mountain mining district before 1967 included 4.6 t (150,000 oz) of gold, 65.3 t (2.1 million oz) of silver, 13.6 kt (15,000 st) of copper, 4.5 kt (5,000 st) of lead, and 13.6 kt (15,000 st) of zinc (Roberts and Arnold, 1965). At the time of Duval's 1967 start-up, ore reserves at Copper Canyon included 12.6 Mt (14 million st) in the East ore body and another 3.6 Mt (4 million st) in the West ore body. Both had average grades of 0.79% copper, 16.1 g/t (0.47 oz per st) silver and 0.86 g/t (0.025 oz per st) gold (Blake et al., 1984). The Tomboy-Minnie deposits collectively contained 3.5 Mt (3.9 million st) of ore grading 3.09 g/t (0.09 oz per st) gold and 9.6 g/t (0.28 oz per st) silver. The Northeast Extension reserves were identified as 1.4 Mt (1.5 million st) of ore with an average grade of 2.9 g/t (0.085 oz per st) gold and 5.1 g/t (0.15 oz per st) silver. Initial total reserve estimates for the Fortitude deposit were 14.5 Mt (16 million st) grading 5.14 g/t (0.15 oz per st) gold and 29.8 g/t (0.87 oz per st) silver (Jackson, 1982). Therefore, before large scale mining, the Copper Canyon area contained a minimum of 102.6 t (3.3 million oz) of gold and 541 t (17.4 million oz) of silver as mineable reserves. The Fortitude deposit reserves as of Jan. 1, 1985 were 10.6 Mt (11 million st) grading at 5.28 g/t (0.154 oz per st) gold and 28.5 g/t (0.83 oz per st) silver. Currently, the company is producing 6.8 t (220,000 oz) of gold and significant amounts of

Jan 7, 1986

By J. Finch, J. Yianatos, N. W. Johnson, R. Espinosa-Gomez

A method of determining the maximum concentrate solids rate or carrying capacity is described for given column operating conditions. The method requires alteration of the feed rate of solids to the column at a fixed residence time by manipulation of the % solid of the column feed pulp or the bias rate. The carrying capacity was measured for five streams at the Mt. Isa Mines Pb/Zn and Cu concentrators. As a potential scale up parameter, the mass of solids per unit time per unit column cross section (Ca, g/min.cm2) is explored. Ca is tentatively related linearly to the 80% passing size (d80) of the concentrate. For d80 < 15 pm Ca was less than 4 g/min.cm2 (8 lb/min.ft2) which makes it a critical consideration in scale-up for such fine feeds. An alternative carrying capacity parameter, mass of solids per unit volume of gas Cg was not as useful as Ca since Cg varied inversely with gas rate.

Jan 1, 1988

By J-P Franzidis

A study was undertaken to determine the effect of five industrial frothers on the relationship between entrainment recovery and water recovery. The experiments were conducted using a Batequip rectangular 60 liter steel cell which was run in parallel with the first cell in a copper/silver prefloat rougher circuit. The concentration of the frothers ranged from 2.5 g/ton to 8.4 g/ton. The experiments were repeated at three different froth depths. A power function was found to fit this relationship with no discontinuity in the relationship caused by the changes in frother type and concentration and froth depth.

Jan 1, 2000

By G. S. Walker, W. van Bronswijk

The accurate determination of free cyanide and metal cyanide complexes is of increasing interest to Carbon-in-Pulp (CIP) gold extraction operations as cyanide consumption is an important economic factor in the processing of complex ores. Characteristically complex ores are high cyanide consumers and these ores usually contain a range of sulfide minerals e.g. covellite (CuS), digenite (Cu1,S), chalcocite (Cu, S), Pyrrhotite (Feci-x)S), marcasite (FeS2), arsenic and antimony sulfides and zinc sulfide. These minerals can consume significant quantities of cyanide, e.g. various copper(!) cyanide complexes such as CuCN, Cu(CN)2·, Cu(CN)/ and Cu(CN)/ can form, depending on the cyanide to copper ratio in solution. Measurement of free and complexed cyanide is thus a prerequisite for efficient cyanide dosage control. Free cyanide can be determined potentiometrically or amperometrically but the methods are prone to error and are laborious [1]. Metal cyanide complexes can be determined by reverse phase HPLC, however, the species ofa single cation (e.g. Cu(CN)i-, Cu(CN);2 and Cu(CN)/) elute as one species [2]. Additionally the high amount of suspended matter in CIP leach solutions tends to introduce errors. Raman spectroscopy has the potential to be used as an in­line or on-line method of analysis. Free cyanide and the cyanides of gold, silver, copper and iron have been studied in aqueous alkaline solutions with a view to developing such a method. Spectra were obtained with a Bruker RFSl00 FT-Raman spectrometer using a 1064 nm laser delivering 800 mW of power at the sample. Depending on the solution, 250, 1000 or 4000 scans at 4 cm-1 resolution were accumulated. The cyanide stretches observed were relatively narrow (10 15 cm-1 full width at half height) and well resolved (Figure 1 ).

Jan 1, 1995

By Louis M. Schlesinger, Joel F. Hutwelker

A pilot plant was constructed to attrition scrub and deslime North Carolina phosphate ore. A 15.2 cm (6 in.) diameter pilot-scale Linatex Hydrosizer was used for desliming. Four tests were run at a throughput of 2.9 t/hr/m2 (0.30 ton/hr/f t2) and a teeter bed density of 1.12-1.13 g/cm3. The teeter water flow rate was varied to create an upflow velocity ranging from 0.25-0.47 cm/s. (This latter value is equal to the terminal settling velocity of a 200 mesh (74 micron) particle having a density of 2.7 g/cm3.) The recovery of +200 mesh oversize to the sizer underflow was 95%-96%, regardless of the teeter flow rate. The recovery of -200 mesh undersize to the sizer overflow increased from about 90% at 0.25 cm/sec to 100% at 0.47 cm/sec. A metallurgical balance that was conducted for one of the tests showed that less than 2% of the phosphatic value was lost to the overflow. A fifth test was run using an upflow velocity of 0.47 cm/sec; however, the teeter bed density was maintained at 1.05 cm/sec. This change had the effect of reducing the cut point (d50), while maintaining complete slime removal.

Jan 1, 1993

By Andreas G. Mueller

Gold-silver and copper-gold deposits with high-temperature wallrock alteration of Ca-Mg-Fe-Al silicates occur in ca. 3.0 Ga old greenstone belts of the Yilgarn Block, Western Australia. The deposits are hosted by amphibolite-facies komatiites, basalts and BIF, controlled by ductile shear zones and located within the contact-metamorphic aureoles of composite granodiorite-granite batholiths. Disseminated sulfides and free gold form foliation-parallel ore zones in a gangue of anhydrous silicates (garnet, clinopyroxene, olivine), hydrous silicates (amphibole, biotite) and calcite. The deposits may be classified on the basis of their dominant economic metals as: 1(1) Gold skarns with substantial silver as by-product, represented by the Marvel Loch mine (production: 2,200,000 t at 3.5 g/t Au), and characterized by the mineral association pyrrhotite, loellingite, arsenopyrite, galena, gold and magnetite. (2) Gold skarns with low silver contents, represented by the Nevoria mine (production:1,850,000 t at 3.4 g/t Au), and characterized by the mineral association pyrrhotite, chalcopyrite, arsenopyrite, gold, Bi-tellurides and magnetite. (3) Copper-gold skarns, represented by the Elverdton mine (production: 850,000 t at .7 5 Cu, 0.81 g/t Au), and characterized by the mineral association magnetite, chalcopyrite, pyrrhotite, pyrite, gold and Ag-Bi tellurides.

Jan 1, 1990

By D. Sotiropoulos

In the recent years, the mining sector is looking for specific indices, taking into account economic, environmental and social criteria, in order to meet the sustainable development requirements. In the present study traditional and environmental sustainable indices used in the Agios Dimitrios Power Plant, Northern Greece, are compared to those of ESKOM Company, South Africa, taken as reference. These indices are referred to, a) the solid fuels consumption, such as lignite, and, b) the emissions to the atmosphere, such as SO2, NO2 and dust. Last five years data show a remarkable improvement of all indices in the Agios Dimitrios Power Plant. Although the lignite used in this Power Plant is by far of poorer quality than that used by ESKOM - net calorific value 1272 and 4638 kcal/kg, respectively - dust emissions in both Power Plants are similar (52 and 31 g/kWh, respectively). The SO2 and NO2 emissions are very low in Agios Dimitrios compared to those of South Africa, respectively 1.92 and 7.91, and 1.74 and 3.61 g/kWh. This is mainly due to natural desulfurization (by the lignite?s calcium content) and the low NOx burners.

Jan 1, 2003

By B. J. I. Adair, K. R. Weller

The bulk scavenger tailing from Aberfoyle&apos;s Hellyer Concentrator, which represents about one-third of the total tailing, runs about 2.5-3 g/t gold and 140-200 g/f silver. Both values are very finely disseminated, particularly the gold. Aberfoyle wished to establish whether ultra-fine grinding followed directly by cyanide leaching could be applied to the tailing for recovery of gold and silver. The work comprised the determination of gold and silver recoveries as a function of grind size and leaching time, the specific energy consumption required for continuous wet grinding of the tailing to the specified product size in a small horizontal continuous stirred mill, and an estimate of the rate of media consumption. The maximum extraction of gold, 23.3% was obtained from the finest product, with the highest cyanide concentration and longest leach time, confirming earlier work. A similar result was obtained for silver. The specific energy requirement for milling rose from 27.7 kWh/t for a P80 of 8.1 um to 219 kWh/t for a Pea of 2.3 um. The consumption of ceramic media was about 1.6 g/kg of final product. An extensive QEM&apos;SEM Trace Mineral Search on the unground tailing failed to identify any gold-bearing particles but did identify silver-bearing particles which displayed a variety of textural associations and liberation characteristics.

Jan 1, 1998

By A. Chieregati

"A heterogeneity study was conducted at Erdenet Copper Mine in Erdenet, Mongolia, to determine the fundamental sampling error associated with blasthole sampling. Using a custom-designed radial sector sampler , six large blasthole samples were collected to perform the test developed by Pierre Gy to estimate the constitutional heterogeneity factor for several types of ore. The study investigated the entire sampling protocol, from mine site to analytical laboratory, and quantified the contributions to total fundamental sampling error (FSE) of the various stages of sampling and sample preparation. The total relative standard deviation (RSD) of sampling error was determined to be ±32 percent and ±49 percent for copper and molybdenum, respectively. A classical statistical analysis conducted on the data confirmed the level of RSD native to the mine. About 95 percent of the sampling error was produced from one step in the sampling protocol, that of reducing the 1-kg (2.2-lb) sample to a 50-g (0.1-lb) sample for chemical analysis. To minimize the FSE, the use of a slightly different protocol is recommended. The 1-kg sample (2.2-lb) should first be crushed to a 0.2 cm (0.08 in.) particle size. A 250-g (0.5-lb) split should then be taken and finely ground, from which a 50-g (0.1-lb) split should then be taken. These simple steps will reduce the FSE to ±7.9 percent and ±17 percent for copper and molybdenum, respectively. The modified sampling protocol effectively reduces to zero the probability of misclassification of low-grade ore. IntroductionThe Erdenet copper-molybdenumopenpit mine is located in Erdenet in Orkhon province, Mongolia. The Erdenetiin-Ovoo copper-molybdenum deposit is 1.5 to 3 km (0.9 to 1.8 miles) wide and about 25 km (16 miles) long (Erdenet Mining Corp., 2016). Chalcopyrite, bornite covellite and chalcocite are among the major minerals found in the deposit. The mine processes approximately 26 Mt/a (29 million stpy) of ore."

Jan 11, 2017

By A. Khoshnevisan

Molybdenum resources around the world are mainly associated with porphyry copper ores. Molybdenite (MoS2) is the main Mo mineral that has rhenium in its crystal lattice. This paper investigates the recovery of rhenium and molybdenum from a molybdenite concentrate using a hydrometallurgical treatment method. The molybdenite concentrate is leached with nitric acid and subjected to solvent extraction to recover and separate rhenium and molybdenum from the leached liquor. The effects of pH and the chemicals? concentration on extraction characteristics of Mo and Re are investigated to define the best condition for selective extraction of these metals. The tested leached liquor contained of 8.2 g/L Mo, 1.5 g/L Fe, 0.5 g/L Cu and 60 ppm Re. A good extraction efficiency, up to 94%, was achieved for Re from the leached liquor, where the extraction of other metals, such as Fe and Cu, was negligible. In particular, at very low pH, rhenium is extracted selectively from solution, while the extraction of molybdenum is less than 10%. After removing rhenium, molybdenum was extracted with LIX984N. The selectivity of molybdenum extraction over other metals in solution was significant. Nearly 94% of molybdenum can be extracted from the solution at a pH value of -0.12, without any interference from other metal ions such as copper and iron.

Draglines are one of the largest land-based equipment moving across the earth. They typically ranging in size from2.5 to 6.5 kt (2,750 to 7,100 st). Their size, though, makes them difficult to maintain. Recently, G&S Engineering Services used an Enerpac synchronous hydraulic lifting system to lift a 3.2-t(3,500-st) dragline at the Curragh Mine in Queensland, Australia’s Bowen Basin.

Jan 1, 2004

By P. Karimi, A. Akcil, A. Amini, H. Abdollahi

"In this study, cyanidation experiments were carried out on a massive sulfide ore sample, whose grades of the precious metals gold and silver were found to be 4.2 g/t and 311.9 g/t, respectively. Other elements in the ore were arsenic with average grade of 1,751.6 g/t; antimony, 619.7 g/t; zinc, 781.4 g/t; copper, 454.6 g/t; lead, 101.8 g/t; mercury, 324.3 g/t; and bismuth, 0.6 g/t. The first objective of test work focused on determining the best conditions for extracting gold from the ore, ground to less than 75 microns. Different influential parameters such as particle size fraction, cyanide concentration, pH and leaching time were investigated. The optimum parameters were found to be particle size fraction of less than 25 microns, cyanide concentration of 2,000 mg/L, pH of 11 and cyanidation time of 24 hours. In these optimum conditions, the highest gold and silver leaching efficiencies were 76.88 and 59.63 percent, respectively. In addition, roasting as pretreatment at a temperature of 700°C over a period of two hours was evaluated. Cyanidation testing of the roasted ore in the optimum conditions found increases in the gold and silver leaching efficiencies to 85.97 and 67.57 percent, respectively. Moreover, antimony extraction rose to 43.10 percent, from 23.73 percent without roasting.IntroductionGenerally, gold ores can be classified as “free milling” or “refractory,” depending on how they respond to cyanide leaching (La Brooy et al., 1994; Celep et al., 2009). While high gold recoveries (> 90 percent) from free-milling gold ores can be readily achieved, refractory gold ores are often characterized by low gold recoveries (50-80 percent) within a conventional cyanide leaching process (Rubisov et al., 1996; Adams, 2005; Celep et al., 2009). The presence of preg-robber components such as carbonaceous matter, sulfide minerals and aluminosilicate can further decrease gold and silver dissolution rates and recoveries. With the negative effect of the presence of preg-robbers, gold dissolution is considerably diminished in solutions with high cyanide concentration (Tan et al., 2005). In such refractory gold ores, gold particles may sometimes be occluded or included in the sulfide minerals, and pretreatment becomes necessary to break down the mineral structure to liberate gold for subsequent recovery. These ores are usually pretreated with an oxidizing process, after which gold and silver can be recovered by a standard cyanidation process. The main oxidizing processes available for refractory gold ores are roasting (LaBrooy et al., 1994; Angelidis and Kydros, 1995; Li et al., 2009; Karimi et al., 2010; Kyle et al., 2012), biological oxidation (Lawrence and Bruynesteyn, 1983; Komnitsas and Pooley, 1990, 1991; Deng and Liao, 2002; Ciftci and Akcil, 2010, 2013) and pressure oxidation (Mason, 1990; Murthy et al., 1992; Qingcui et al., 2009). In addition, ultra-fine grinding is being used in some plants for the liberation of gold and silver particles that are interlocked in the sulfide or silicate minerals (Celep et al., 2011)"

Jan 1, 2015

By B. H. Kim

The use of a 25.4 mm diameter water-only hydrocyclone for separating fine-size (-150 µm), high-density (>4.5 g/cc) particles from similar-size, lower-density (2.65 glee) particles was investigated. It was found that a single-stage separation recovered nearly 90% of the heavy minerals (7.0 g/cc), while achieving a volume reduction of about 70%. The use of dense medium in a single stage improved volume reductions to 90% at similar heavy mineral recoveries. In addition, computer simulations of a three-stage circuit were performed, which showed that recoveries approaching 97% with volume reductions of 95% were possible.

Jan 1, 1998

By Robert M. St. Louis, Janet M. Edgecombe

The Mercur Mine&apos;s autoclave is an alkaline pressure oxidation system, operating at 220°(428°F) and 3061 kPa (444 psig) (total). Up to 725 metric tons (800 st) of ore, grading 2.02 g per mt (0.059 oz/st) Au and containing 1 to 2% sulfide sulfur, are treated daily in this circuit. A mineralogical and chemical investigation was initiated to locate the cause of lower than anticipated gold recoveries. Greater than 70 wt. % of reduced sulfur is oxidized; therefore, the autoclave is operating at specification. The low recoveries were attributed to increased concentrations of organic carbon in refractory ore. Activated carbon concentration in the attached CIL circuit was increased from 15 to 40 g per 1, leach temperature held below 21°C (70°F), and carbon flow changed to counterflow. These measures resulted in gold recovery improving from 75 to 82%.

Jan 1, 1990

By N. Parra, M. N. Herrera, T. Vargas, B. Escobar, C. Gonzalez

originally published May 1998, Vol. 15, No. 2, pp. 15-1 9 Discussion by G. Rossi, University of Cagliari Reply by T. Vargas, University of Chili

Jan 1, 2000

By V. B. Menon, L. D. Michaels, M. E. Mullins

Introduction The cleaning of coal to remove sulfur and inorganic impurities has, in recent years, become an essential operation in any coal utilization process. US coals, on average, contain about 3% total sulfur, of which more than 60% is generally inorganic or pyritic sulfur. Existing technologies such as heavy media cycloning, selective agglomeration, and electrostatic separation have the potential for removing about 80% of the pyrite from coal (Khoury, 1981). In view of the large volumes of coal being used today, there exists a need for the development of new processes to further improve the separation of pyrite from coal. Colloidal particles in suspension possess surface charges that are related to the chemical composition of the particles. In aqueous suspensions, the surface charge is a function of pH. For coal-pyrite mixtures, it is possible, by an adjustment of the pH (also a function of polar molecules), to reach a state where the charge on coal is very close to zero, but the charge on pyrite is still substantial. The introduction of collector particles with a charge opposite that of the pyrite should then result in the selective flocculation of pyrite and collec¬tor. These flocculates should settle out much faster than coal and could be removed from the bottom of the settling column. This note discusses our preliminary observations of the effect of clay-based collector particles, especially bentonite, on the settling behavior of coal and pyrite. These experiments were conducted with a view to developing a new electrokinetic flocculation process for separation of coal-pyrite mixtures. Experimental The objective of the experiments was to determine the effects of clay-based additives on the settling rates of clean coal, pyrite, and coal-pyrite mixtures. Illinois Colchester (No. 2) coal was obtained after processing in an existing coal cleaning facility. This coal was reported to contain less than 0.3% pyritic sulfur, which was in the form of a fine powder with an average particle size of 5µm. Pyrite (of average particle size 5 µm) was also obtained from the same facility and reportedly contained less than 0.1% coal. The zeta potentials of coal and pyrite in water were measured as a function of pH using a zetameter. The pH was adjusted by adding sodium hydroxide or hydrochloric acid to the suspensions. Sedimentation experiments were conducted in 50-mL glass cylinders, and settling rates were measured by noting the height of clear liquid above the settling front as a function of time. Settling behavior was also charted by periodically taking photographs of the system. The following systems were investigated: •a slurry consisting of 2.5 g of clean Illinois Colchester coal in 50 mL of water of pH 6.0, with and without the addition of 0.5 g of bentonite (Fisher, purified grade) ; • a slurry consisting of 2.5 g of pyrite in 50 mL water of pH 6.0, with and without the addition of 0.5 g of bentonite ; and • a slurry consisting of a mixture of 1.25 g clean coal and 1.25 g pyrite in 50 mL water of pH 6.0, with and without the addition of 0.5 g of bentonite. In addition, experiments were conducted with kaolinite (Bath, South Carolina) and alumina (Fisher, purified grade) as collector materials to verify the observed effects. Results and discussion Figure 1 shows the zeta potential values for the Illinois coal and pyrite as a function of pH. The zeta potentials of coal and pyrite decrease with increasing pH, changing from an initially positive value at low pH to a negative value at high pH. The point of zero charge (PZC) for the coal is at a pH of 6.2, while the PZC for pyrite occurs at a pH of 10.3.

Jan 1, 1988

By Fathi Habashi

METALS COMMONLY leached may come from one of the following sources: 1) Occurring in nature in the native state, e.g., gold, silver, platinum metals, copper. 2) Produced as by-product of a metallurgical process, e.g., gold, silver, platinum metals, selenium. 3) Produced by reduction of an oxide ore, e.g., copper, nickel. GOLD AND SILVER Cyanidation According to Mellor (1923) the alchemists of the eighteenth century knew that gold was soluble in aqueous solution of potassium cyanide, and the early jewelers probably utilized this fact in gilding. He also mentions that Scheele in 1783 and later Bagration in 1843 noted the solvent action of aqueous solutions of alkali cyanides on gold. Yet it was not until 1887 that the value of this observation was recognized, when MacArthur and Forrest (1887) made use of it in leaching gold from its ores. Since then, the process has spread to every gold and silver mining district in the world.

Jan 1, 1970

By E. Wm. Heinrich

Glass sands fall into that extremely select category of mineral deposits that requires a fantastic degree of natural concentration (95% +) of a single component before they can even be considered as ores suitable for further refinement. The combination of geological factors that has produced such an environment of "mineralogical single-mindedness" has appeared only rarely. Factors that require consideration in the economic evaluation of a potential glass-sand deposit can be grouped into three categories: 1) geological (e.g., size, shape, structure); 2) petrographic (e.g., grain size, accessory minerals, nature of iron contaminants); 3) geographic - political (e.g., proximity to consumer, size of market, zoning restrictions). Of particular con ern is the nature of the iron contamination. Iron-bearing species appear as 1) accessory mineral grains, either detrital or authigenic, 2) cements, 3) grain coatings, and 4) quartz-grain inclusions. Selection of the cheapest form of beneficiation requires precise information on the mineralogy of the iron species and their detailed petrography. Because of the antipathetic relationships, in many parts of the United States and Canada between the geographical locations of glass plants and the geological distribution of glass sands, the selection of ac suitable source for a particular glass plant may be dichotomous. One may seek either an exceedingly pure "classic" type glass sand that is relatively distant and whose cost includes a major transportation charge or a relatively proximate and somewhat "impure" source whose cost includes a substantial fee for beneficiation.

Jan 1, 1974

By Mahesh C. Jha, Frank J. Smit

Column flotation and selective agglomeration are promising advanced physical cleaning processes that can be developed for preparing premium quality coal-water slurry fuel from coal fines. Such a fuel could either replace oil or natural gas in certain industrial and utility boilers or else could be burned in advanced combustors currently under development. These applications require coal cleaned to sulfur and ash contents below 258 g/GJ and 430 - 860 g/GJ (0.6 lb/million Btu and 1 - 2 lb/million Btu), respectively, to meet clean air emission requirements. A highly loaded slurry must be produced in order to avoid derating the boilers, and the total cost of the fuel must be below $2.37/GJ ($2.50/million Btu) in order to penetrate the market by the turn of the century. Achieving these technical and economic goals depend critically on the selection of suitable feed coals. This paper describes the identification and evaluation of 18 coals and final selection of six coals from different regions of the United States as potential feedstock for preparation of premium fuel. The work is being performed under a cost-shared DOE project.

Jan 1, 1995

By K. Park

Coupled influence of sulphidizers and xanthates on the electrokinetic property and flotation behavior of malachite was systematically examined via electrophoretic mobility measurements and microflotation tests. The electrophoretic mobility results for the malachite before and after surface modification with different types (Na2S and (NH4)2S) and amounts (1×10-7?1×10-4 moles/g) of sulphidizers, showed that the isoelectric points (IEP) of malachite differed at the low dosage level depending on the type of the sulphidizer, which ranged between 5 and 9.5. However, the IEP values were observed to be shifted from high to low pH values with increasing amount of sulphidizers and finally ranged between 3 and 4, indicating that the sulphidizers played a different role on changing the surface property of malachite. In order to further understand the effect of sulphidizers on the electrokinetic property and flotation behavior of malachite in a sulphidizer+xanthate system, flotation tests were carried out with malachite sulphidized with Na2S and (NH4)2S (adsorptive amount=8×10-5 moles/g) at different bulk xanthate concentrations. The results showed that the flotability of (NH4)2S-treated malachite was greater than Na2S-treated one over the entire xanthate concentration investigated, indicating that the type of sulphidizers played important role in malachite flotation; however, no obvious correlation between electrokinetic property and flotability of malachite was observed.

Feb 27, 2013