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By K. Takahashi

The methyl stretching vibrations for a series of normal and iso xanthates were determined using an infrared spectrophotometer, and the results were compared with the published information on the corresponding alcohols. The frequency shifts of methyl stretching vibrations may be attributed to the inductive effect of polar groups. The hydrogens in the shorter chain alkyl groups are somewhat polar and can be expected to interact with water molecules by hydrogen bonding. Contact angle, surface tension lowering, and solubility appear to be related to the induced polarity of the alkyl groups.

Jan 1, 1968

By Guy T. Jr. McBride

World food production is not matching the exploding increase in world-population. Since arable land is limited, ways are being sought to increase yields by turning_ to fertilizer. Phosphate, an essential fertilizer ingredient, is undergoing a period of growth unparalleled in history. The United States is at once the world's largest producer of food and of phosphate rock. Both food and fertilizer output in the United States are being enlarged to meet the demands posed by the threatening world food crisis. New phosphate deposits are being developed in Florida and North Carolina to assure the United States' continued preeminence in this vital resource. Large fertilizer complexes are also being constructed with an eye to both domestic and foreign markets. The growth of the U. S. phosphate industry has been characterized by developing technology in the processing and distribution of phosphate raw materials. The trend from low analysis to high analysis materials accompanied by a change from a large number of small fertilizer plants oriented toward the marketplace to a fewer number of large plants located at the raw material source is illustrative. These changes in technology and distribution have also been important factors in the development of new mining areas in Florida, the Western States and North Carolina.

Jan 1, 1966

By Thomas E. Garnar

Mineral sand products have been produced from beach sand heavy minerals for nearly a century. These minerals have specific gravities of 3.0 or higher and are extremely resistant to weathering. The most common economic heavy minerals are ilmenite, leucoxene, rutile, zircon, stauarolite, monazite, garnet, and the aluminum-silicate minerals. Of these, ilmenite is the most abundant. In some cases a deposit may contain only one or two economic heavy minerals whereas in others five or six may be present. Heavy mineral mining began in the late eighteen hundreds when monazite was mined for its thorium content. This was followed by mining activity on the North Florida coast just south of Jacksonville Beach - where ilmenite and zircon were first separated during the WW I (1,8). Modern heavy mineral mining began in Eastern Australia during the late nineteen-thirties and spread to the United States during World War II. It is the purpose of this paper to review the commercial heavy mineral products, their sources, and their uses.

Jan 1, 1985

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

"The Analysis of Retreat Mining Pillar Stability (ARMPS) program was developed by the National Institute for Occupational Safety and Health (NIOSH) to help the U.S. coal mining industry to design safe retreat room and pillar panels. ARMPS calculates the magnitude of the in-situ and mining-induced loads by using geometrical computations and empirical rules. In particular, the program uses the “abutment angle” concept in calculating the magnitude of the abutment load on pillars adjacent to a gob.In this paper, stress measurements from U.S. and Australian mines with different overburden geologies with varying hard rock percentages were back analyzed. The results of the analyses indicated that for depths less than 650 ft, the ARMPS empirical derivation of a 21° abutment angle was supported by the case histories; however, at depths greater than 650 ft, the abutment angle was found to be significantly less than 21°. In this paper, a new equation employing the panel width to overburden depth ratio is constructed for the calculation of accurate abutment angles for deeper mining cases.The new abutment angle equation was tested using both ARMPS2010 and LaModel for the entire case history database of ARMPS2010. The new abutment angle equation to estimate the magnitude of the mining-induced loads used together with the LaModel program was found to give good classification accuracies compared to ARMPS2010 for deep cover cases.INTRODUCTIONIn the early 1990s, the Analysis of Longwall Pillar Stability (ALPS) was introduced by Mark (1990) as a chain pillar design software and was generally accepted and used by the U.S. coal mining industry. Following the success of ALPS, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) program for designing retreat mining pillars using a similar approach as ALPS (Mark and Chase, 1997). The Australian mining industry also recognized the success of ALPS, and Colwell, Frith, and Mark (1999) calibrated the program to Australian conditions. The ALPS and ARMPS programs draw their strengths from the large databases that are used to calibrate them (Mark, 2009). However, following the Crandall Canyon Mine collapse in 2007, NIOSH had to reconsider the pillar design criteria used in deep-cover retreat mining (Mark, 2010). The ARMPS overburden load prediction algorithm was improved to more accurately predict the loading of narrow panels with high overburden depths by implementing the pressure arch concept, and this new version is called ARMPS2010."

Jan 1, 2019

By W. C. Morgan

The Federal Surface Mining Law enacted in 1978 is now being translated into regulations which must be followed in future coal mining. Substantial change to mining procedures is required to maintain stricter air and water control while mining and to return the mined out land to equal or better use than before. These changes do not come easy; and in most cases, they create a substantial additional cost. Based on field observations and mining studies, Caterpillar has been making suggestions on the use of their types of equipment to conform to the new standards at the least cost. This paper will be a brief review of some of the overburden stripping and reclamation methods suggested by Caterpillar. A production and cost comparison of dragline, scraper/dozer and truck, loader systems in 20 to 150 ft. Overburden conditions illustrates potential cost savings using mobile equipment.

Jan 1, 1980

By Robert L. Somerville

As more phosphate deposits are discovered and new mines are developed, the necessity of critically evaluating various rocks for phosphoric acid production recurs ever more frequently. This need is accentuated by the increasing number of, fertilizer plants being constructed on deep water at various locations to serve world markets and to take advantage of locally-available sulfur or sulfur-bearing ores. A major proportion of all phosphate rock mined is consumed by wet-process plants and this acid in turn is consumed, usually captively, in the manufacture of fertilizer. It is apparent, therefore, that the first question asked by the developer of a new deposit is - How will it work in a wet-process plant? and conversely, when developing a deep water, project, one of the early studies required is - Which of the available rocks shall we use? What is the correct price to pay? How do we compare one rock with another? The economic importance of even minor aspects of these questions

Jan 1, 1966

By Thomas D. Henderson, Wayne R. Hopkins, Arthur Lynch

Brought on stream in 1975, the 10,000-tpd copper oxide processing plant of Anamax Mining Co. at Twin Buttes, Ariz., is now at design capacity. Based on acid agitation leaching, solvent ion exchange, and electrowinning, the Anamax plant bridges an ore to metal gap by producing mine site copper cathodes from lime-bearing oxidized feed. Until recently, such material was considered submarginal because copper recovery was uneconomic using conventional methods. The new Twin Buttes oxide plant, the largest of its kind in North America, is one stage in the expansion of Anamax, a joint venture owned equally by The Anaconda Co. and Amax Inc. Other phases include an accelerated stripping program in the pit, an expansion of the sulfide concentrator from 32,000 to 40,000 tpd, and construction of a 57 MW diesel generating station for the oxide plant. Once the basic oxide flowsheet had been selected, Anamax awarded a contract to a joint venture of Arthur G. McKee and Co.' (Western Knapp Engineering Div.) and Davy Powergas' (a member of the Davy International group). The contract covered process design, engi¬neering, procurement, and construction. The property, identified by Banner Mining Co. in the 1950s and developed by Anaconda, was put into production in 1965 as an open pit mine equipped with a 32,000-tpd sulfide flotation concentrator. The Twin Buttes mine has since become one of the largest earth¬moving operations in the world, excavating some 10 tons of waste per ton of ore. At the original sulfide mill feed rate, the mining fleet moved 375,000 tpd of material (225,000 tpd alluvium and 150,000 tpd of rock). The most significant copper mineral in the ore body is chalcopyrite, which is scattered in pockets of weak mineralization. Cap oxidation has progressed to depths of several hundred feet, principally as chryso¬colla. Present estimates place minable reserves at 347 million tons of sulfide ore grading 0.6% copper and 55 million tons of oxide ore grading 0.9% recoverable copper from 1.25% total copper. The nonre¬coverable fraction is composed of native copper and insoluble copper oxides and sulfides. The oxide ore has been partially stripped, and 25 million tons having a copper content greater than 0.6% have been stockpiled. The stripping of oxide ore to expose sulfides will continue for several years. The copper oxide studies launched by Banner Mining in the 1950s were continued by Anaconda. Process routes investigated included: acid leaching; cyanide leaching; segregation roasting; chlorination, oxidation, and reduction; flotation in combination with leaching; cal¬cite flotation plus leaching; copper oxide flotation; and sulfidization flotation. By late 1968 test work favored a low temperature roast followed by an ammonia-ammonium carbonate leach of the ore. During this period, acid leaching had been discounted because of excessive con¬sumption of expensive acid by the high limestone ore matrix. Restric¬tive SO2 emission controls in 1971 suggested an increase in acid production and availability at a lower cost. Research then switched to acid leaching, resulting in development of the present flowsheet. Mechanically agitated leaching in a cascaded series of tanks was selected in preference to vat leaching after piloting a stirred-tank system. Preliminary testing of vat leaching indicated that inter-vat solution clarification problems could be caused by gas evolution from leaching of high limestone ore. Lack of research time for resolution of this problem, plus success of the pilot agitation system, dictated the final flowsheet selection. In 1973 Anamax decided to proceed with agitation acid leaching to produce liquor for a solvent ion exchange (SIX) and electrowinning (EW) installation. At that time, such systems were proving their eco¬nomic viability, and they allowed direct recovery of high quality cath odes while bypassing the smelting and refining of lower grade cement copper precipitates. The Anamax oxide plant processes a grade of about 1% recovera¬ble copper from stockpiled sources and new mine production. Unit operations include size reduction, agitation leaching, countercurrent decantation washing of residue, pH adjustment of solution, clarifica¬tion, solvent ion exchange, and electrowinning of copper, with a mini¬mum of environmental impact. The feed to the primary crushers is composed of: 1) -10-in. stockpiled oxide ore, which has passed through the mine crusher and has been grade-segregated to allow blending for a consistent feed material. 2) -4-ft ROM oxide ore stockpiled in the early mining operation. 3) -10-in. ore currently being mined. Ore is delivered in 100-ton mine trucks to a dump pocket, where an apron feeder transfers the ore to an inclined 6-in. fixed grizzly. The grizzly oversize passes through a 48 x 60-in. primary jaw crusher, where it is reduced to -6 in. The fixed grizzly undersize and the crusher product are conveyed to a 4-in. vibrating grizzly. Oversize from this grizzly reports to a 7-ft Standard cone crusher, where it is reduced to -1 in. The crusher product and the vibrating grizzly undersize are conveyed to two 8 x 20-ft double-deck vibrating screens that operate in closed circuit with two 7-ft short head cone crushers. Screen undersize (-1/2 in.) is delivered to a covered conical fine ore stockpile with a designed live capacity of 15,000 st. Beneath the stockpile are ten Pioche belt feeders feeding two conveyor belts that supply the milling circuits. Each group of five feeders has two variable speed and three fixed speed drives. Grinding to 95% -48 mesh is done in two parallel lines, each consisting of an 111/1 x 18S -ft rod mill and a 12' x 30-ft ball mill in open circuit. Water is added to maintain a pulp density of 72% solids in the mills. Further water additions are made to the ball mill discharge sumps to create a 60% solids slurry for transfer to the leaching reactors. The nominal grinding rate is 440 tph solids. A splitter box on the product from the mills separates the incoming slurry into two streams, one for the leaching reactors and the other for the pH adjustment reactors. This split is determined by the lime content of the feed material and the pH of solution entering the pH adjustment reactors. Five 30-ft diam by 31-ft high mechanically agitated, rubber-lined leach tanks are arranged in cascade, with gravity transfer from one to another by enclosed launders. Concentrated sulfuric acid (93.2%) from storage is added to the pulp principally at the first leach tank, and the reacted slurry leaving the cascade after a 5-hr residence time at 50% solids is pumped to countercurrent decantation (CCD) wash thickeners.' At this point, most of the soluble copper has been leached from the ore. Up to 250 lb of sulfuric acid is consumed per ton of ore, the acid being supplied in tank cars from a nearby smelter. Leached slurry at 50% solids is washed countercurrently with return raffinate from the SIX plant in a series of four Doff-Oliver center caisson, 400-ft diam thickeners. Solids advance from thickeners No. I through No. 4 to the tailings dam. SIX raffinate advances from thickener No. 4 to thickener No. l, where a pregnant solution overflows to pH adjustment. To optimize solvent extraction, the pH of the pregnant solution from CCD thickener No. I is adjusted from 1.5 to 2.5 by adding unreacted ground slurry (split from the grinding circuit) to the preg¬nant overflow of No. I CCD thickener. The resulting slurry is retained for 45 min in three mechanically agitated leach tanks. CO, generated during leach and pH adjustment is ducted to a demister, where entrained acid droplets are removed. The cleaned CO, and the No. 4 CCD underflow are the only effluents from the plant. The underflow of No. 4 thickener mixes with tailings from the sulfide concentrator and is neutralized. Slurry exits the pH adjustment vessels at about 10% solids and

Jan 1, 1985

By X. K. Jiang

Upgrading molybdenite ore can be technically difficult and expensive if talc is a major impurity in the ore. The layered structure of molybdenite and talc decreases the surface specificity of floatation and suppression agents, thereby making molybdenite ore beneficiation inefficient if under some wet processing methods. However, oxidic talc surfaces may be distinct from the sulfidic or oxysulfidic molybdenite surfaces, such distinction related to electron work functions and/or densities of electron donors/acceptors sites. If such differences are present, the upgrading of finely sized molybdenite ore by dry triboelectrostatic methods may hold advantages over traditional wet methods.

Jan 1, 1999

By James J. Scott

To the safety engineer, mine planning engineer, and mine inspector associated with an underground operation, stability of the underground structure is of the utmost importance. In the United States in 1961 rock falls killed 167 miners, crippled many more, and cost the industry a fortune in lost equipment and production. Much research has been carried on to try to find ways to eliminate this loss of life and property. Within the past few years the interest in rock mechanics as a science has been at its highest point, Private industry, government agencies, universities and foundations have launched studies to gain a better basic understanding of the stability of underground openings.

Jan 1, 1966

By Frank W. Schmechel

The subsidence monitoring program at the Old Ben No. 24 Mine was planned in 1976 as part of the on- going demonstration of longwall mining. The construction and installation of the network of surface subsidence monuments was started in Play of 1976 and completed four weeks later after which an initial reference survey was made of all surface monuments. Mining of the first longwall panel started in September of 1976 and was completed in May of 1977. Mining of the second panel started in August of 1977 and is continuing at this time. The Automatic Data Acquisition System (ADAS), which is the subject of this paper, was installed over the second panel on March 29, 1978 near the panel center (Figure 1) and approximately 104 meters (370 feet) in front of the face position at that time. Mine No. 24 is located approximately 5 kilometers (3 miles) Northwest of Benton, Franklin County, Illinois. Coal is being mined from the Illinois (Herrin) No. 6 seam which is at a depth of about 189 meters (620 feet) at the site. The Herrin No. 6 is up to 2.7 meters (9 feet) thick in the mine but a constant thickness of 2.1 meters (7 feet) is being mined by the demonstration longwall with both roof and floor coal being left for stability. Although the seam is flat laying at the mine, it dips north- ward regionally. Topographically, the surface over the demonstration panels is essentially flat and low laying with several areas of standing water. The Rend Lake Dam (United States Army Corps of Engineers) is about 305 meters (1000 feet) north of the demonstration site.

Jan 1, 1979

By G. R. Goldak

All of Canada's older uranium mining areas have been intensively prospected and known outcrops have been repeatedly examined. The only ground in these areas which has not been explored is that which is covered by water or other overburden. The discovery of new mines in these areas will, in the author's opinion, depend on the development of new geophysical or geochemical tools capable of locating uranium deposits which are-buried beneath tens to hundred of feet of overburden. Such tools would be extremely useful in Saskatchewan's Cluff Lake and Wollaston Lake areas. Uranium has recently been discovered in significant quantities in these regions but exploration is made extremely difficult and expensive by an almost uniform overburden cover. The author's experience has been confined to Canada but apparently a subsurface uranium prospecting method could be widely exployed in the western United States and Australia as well. Radon-222 measurements have been used for more than a decade on practically a worldwide basis in an attempt to obtain a useful method for subsurface uranium prospecting. It is well known that ha-222 produced by buried uranium deposits can diffuse through an overburden cover so that it can sometimes be detected at surface. An advantage of this method of prospecting is that Rn-222 can be detected with a simple scintillation, ionization or semiconductor detector. One of the essential weaknesses of the method is that the mean diffusion path length of Rn-222 is limited by its very short (3.8 days) half life. For a deeply buried deposit, vitually all of the Rn-222 atoms will decay before reaching surface and, therefore, such deposits cannot be expected to produce a detectable radon anomaly. Thus, except in unusual cases where groundwater or ground air flow assists the movement of radon towards surface, the depth of exploration is probably limited to ten or twenty feet of overburden cover. The high solubility of Rn-222 in water is also a source of difficulty. In an area where overburden covers bedrock with water saturation of the lower overburden levels, the radon will simply dissolve in the groundwater, and soil gas readings at surface will indicate falsely low

Jan 1, 1974

By J. L. Kohler

The knowledge-based expert systems (KBES) is a potentially powerful technique to the mining engineer in traditional areas of mine design and operations, as well as in the emerging areas of mining automation and robotics. Since KBES is still experimental in many respects, caution must be exercised as this "high tech" tool is developed and applied to mining. The objective of this paper is to demystify expert systems by introducing the elements of KBES, discussing the strengths and weaknesses, and identifying some areas of application in mining, based on the authors' experiences.

Jan 1, 1986

By M. F. Dibble

The Fosfago phosphate rock concentrator is located in the state of Goias, Brazil, 300 km south of Brasilia. The ore is from an enriched carbonatite, which was discovered in 1971 during exploration for pyrochlore niobium ore. The flowsheet was developed in Brazil by personnel from Agrico Chemical Company and Consolidated South America. Engineering was provided by Codemin, a subsidiary of Consolidated, Engineering and construction management was pro- vided by both Agrico and Consolidated personnel who formed the Fosfago team. The concentrator was designed for 500,000 tpy of 37-38% P2Oj dried product. The facilities consist of open pit mining, crushing, stockpiling, and concentration. The product is pipelined seven km to a filtering, flash drying, and railroad loadout station. The plant was successfully started in May 1979.

Jan 1, 1980

By K. Galos

The paper presents increasing importance of mineral secondary and waste raw materials in the minerals management in Poland. Definition of mineral secondary and waste raw materials is proposed, according to current Polish law. On this basis, sources of various types of mineral waste raw materials are analysed. The special attention is paid to the wastes from mining and processing, wastes from power plants, and wastes from steelworks and other smelters. Other types of waste and secondary materials, such as metallic containers, glass containers, batteries, metallic wastes and scraps, are also taken into account. Analysis of sources of such materials is shown together with the main directions of their economic use. Available sources of information on manufacture and use of the mentioned secondary and waste raw materials are also analysed. Finally, for a wide range of waste materials, their current and expected significance as substitutes for some industrial and construction minerals in Poland is given.

Jan 1, 2011

By J. J. Zimmerman

Over-exposure to noise remains a widespread, serious health hazard in the U.S. mining industries. Most other categories of illnesses and injuries associated with mining have improved, with the exception of hearing loss. Mine Safety and Health Administration (MSHA) coal noise data from 2000-2004 show that the continuous mining machine is first among all equipment with 35% of the noise overexposures. Joy Mining Machinery, in collaboration with the National Institute for Occupational Safety and Health (NIOSH), is conducting research to reduce noise generated by continuous mining machines. This paper describes a ?JOY Designed for Noise Reduction? (JOY-DNR; Patent Pending) tail section as a noise control for reducing the noise overexposures of continuous mining machine operators. Underground results show a 45% and 65% noise exposure reduction for the operator when compared to a standard machine. Utilizing this newly developed noise control, along with previously proven controls, will provide continuous mining machine operators an opportunity to be within the MSHA-Permissible Exposure Limit (MSHA-PEL).

Jan 1, 2010

By G. Hwang, S. Park, H. Kim, J. Choi, P. Seo

"In the mineral processing, clay minerals have negative effects on the flotation of valuable minerals. In this study, we investigated how the floatability of malachite (CuCO3·Cu(OH)2) is affected by presence of the clay minerals in the fundamental aspects. Flotation experiments were conducted by using Na2S and PAX (potassium amyl xanthate), and the floatability of malachite was compared between the cases in the presence and absence of clay minerals. The results showed that floatability was greater in the absence of clay minerals than in the presence of clay minerals and decreased with increasing weight percent of kaolinite under same conditions applied. In addition to examine how clay minerals affect malachite floatability, we conducted flotation experiment with changing clay adding order. The results indeed showed that malachite floatability was affected by clay adding sequence. INTRODUCTION The demand of copper is growing as industrial development increases, but mineral resources are limited. Therefore, studies for Cu oxide flotation are being conducted to improve their recovery(1-3). However the mining industry is well aware of the difficulty in treating copper ores in the presence of clay minerals(4). Clay minerals have unique properties and these properties may affect flotation. In the previous studies, the effect of slime coating of clay minerals on flotation was investigated(4-6). For example, Zhao and Peng (2014), Peng and Zhao (2015), and Arnold and Aplan (1986) found that clay slime coating on chalcopyrite, chalcocite and coal surfaces through the electrostatic attraction reduced surface hydrophobicity and then depressed the flotation. Other researchers have also reported clay minerals have impact on pulp rheology. Especially in the case of bentonite, pulp viscosity increased with increasing amount of bentonite and the higher pulp viscosity have deleterious effect on copper recovery(7, 8). However, these studies only focus on slime coatings and pulp rheological behavior. Forbes et al. attempted to decoupling rheology and slime coatings effect on flotation but this study has limitation that individual role of two clay species either attached to mineral or suspended in bulk solution was excluded(9). Hence, in this study, we aimed to investigating the effects of individual clay (suspended and attached) on malachite flotation."

Jan 1, 2016

By P. W. Erickson

Development of the Twin Buttes mine entailed the largest known preproduction stripping job in the history of mining. It linked earthmovers and conveyor belts in a new approach to development of a pit that may reach, or even exceed, dimensions of 1 x 1-1/2 miles at the surface and go as deep as 1800 feet. The orebody to be stripped was covered by an average of 460 feet of alluvial gravels, which presented an initial stripping job of great magnitude. The initial pit layout and the original time schedule established early in 1965 called for the excavation and disposal of about 250,000,000 tons of alluvial overburden before the end of 1969. Allowing about a year for equipment delivery, this schedule left four years to do the job averaging about 62-1/2 million tons per year. In 1964 when The Anaconda Company undertook the development of Twin Buttes, it was recognized that special planning was necessary for the disposal of waste material and tailings. Environ- mental control was a prime consideration and is particularly sensitive in the Twin Buttes area due to the proximity to residential developments. Disposal of the 250,000,000 tons of alluvium, as well as future rock and tailing waste material, presented a problem which Anaconda engineers resolved by designing a plan to spread out much of the alluvium over an extended area, including a massive 11- 1/2 mile dike system which will impound mill tailing in three basins. The tailing will be impounded behind these-dikes to avoid wind exposure and blowing. The dikes also serve as a

Jan 1, 1974

By J. J. Poveromo

"Iron ore for alternate iron processes concerns alternate iron products designed for use in EAF (electric arc furnace) steelmaking processes. These alternate iron products include both DRI (direct reduced iron), HBI (hot briquetted iron) and merchant pig iron. Merchant pig iron can be produced in conventional blast furnaces, mini blast furnaces (including charcoal BF's) or novel processes such as RHF/SAF (rotary hearth furnace/submerged arc furnace) such as Iron Dynamics or RHF such as Mesabi Nugget. These novel processes are coal based and can use iron ore concentrates or recycled steel plant oxides. The blast furnace processes can use conventional feed materials: pellets, sinter, lump ore and coke. However, DRI and HBI processes are predominantly pellet based but require DR grade pellets that are low enough in acidic gangue (SiO2 and Al2O3) to avoid additional expense in the EAF process This paper will focus on the quality and availability of DR grade pellets for NAFTA DRI/HBI plants but also concentrates for fines based processes and feed materials for merchant pig iron production. INTRODUCTION This paper will outline alternative (to the blast furnace) competitive ironmaking processes and process routes. Within each section the types of iron ore to be utilized will be discussed. Before discussing some of the details of these processes, it may be appropriate to review some definitions and commonly used terms within this industry. DefinitionsDirect Reduction: Reduce iron oxide to metallic iron without melting. Unreduced ore compounds remain as undesirable oxides • Direct Reduced Iron: Iron oxide feedstock exits in same form as entered (pellets in, pellets out; lumps in, lumps out) • Hot Briquetted Iron: DRI that has been hot (1200°F, 650°C) briquetted to a high density pillow shaped briquette • Hot Metal: Molten iron in liquid form, above 2500°F, 1370°C • Pig Iron: Solid product of the iron blast furnace • Residuals: Undesirable elements such as copper, nickel, chromium, tin, sulfur, molybdenum and phosphorus • Gangue: Rock minerals in the iron ore such as silica (SiO2), alumina (A2O3), calcia (CaO), magnesia (MgO). These remain in the oxide form in DR processes. • Reduction: Fe2O3 + 3 CO = 2 Fe + 3CO2 2 Fe2O3 + 3 H2 = 2 Fe + 3H2O (Fe2O3 > Fe3O4 > FeO > Fe)"

Jan 1, 2018

By M. G. Ayat

A novel tracing technique employing specifically prepared trace particles has been applied to coal preparation separatory systems. The technique employs individually prepared trace particles representing specific known sizes and densities of coal particles. These trace particles are injected into the feed, to the processing circuit, and are then detected at locations in the coal product and refuse streams. It is anticipated that on-line input of the generated partition data will be linked with computer technology to provide rapid analysis of the processing efficiency, and to allow for process control. This tracing technique can revolutionize many processing operations from the diagnostic viewpoint as well as from the coal processing standpoint.

Jan 1, 1986

By P. R. Dugan

The ultimate objectives and interest in removal of sulfur from high sulfur coal include: (a) reduction of the atmospheric load that con- tributes to the problem of acid precipitation; (b) reduction of ash which contributes to decreased energy transfer in boilers and the problem of ash disposal; and (c) alleviating the social and economic dislocations related to uneven geographical distribution of sulfur in coal reserves. The cost of sulfur removal in relation to the value of refined coal will govern the development of an emerging coal-refining industry. The value of refined lower sulfur coal will be influenced by the national commitment to the environment as well as economic considerations.

Jan 1, 1986