Search Documents

By Paul R. Wieber

This paper describes the U. S. Energy Research and Development Administration's (ERDA) program to develop in situ coal conversion processes, especially gasification, he potential of this technology, the technical directions being explored, and the program strategy to carry the process concepts toward commercialization are emphasized. The geographic location of the U. S. coal resources and a market analysis are described. Commercially successful in situ coal processes could add over a trillion tons of coal to the U. S. coal reserves. The economics of these processes is compared with surface gasification independent studies show this process to be 25Z cheaper than surface gasification for Western coals. The major projects are briefly described. The Linked Vertical Well process is in the lost advanced stage and shows great promise. The last test produced a low-Btu gas averaging 171 Btu per standard cubic foot for average pro¬duction rates of 5.5 million standard cubic feet per day. The other projects are Packed Bed, Longwall Generator, and Steeply Dipping Beds.

Jan 1, 1977

By Roe-Hoan Yoon

INTRODUCTION In 1960, coal accounted for 49% of the world energy consumption, an amount that was reduced to 29% in 1973 because of the availability of cheap oil. At that time, it seemed that oil would remain available for a long time in unlimited quantities. However, this situation changed drastically in 1973/74 when a series of closely spaced price increases from $3 per to $12 per bbl by mid-1974 hit both industrialized and developing countries. Another sharp price increase in 1979/80 to the $34 to $38 per bbl level also hit the world economy hard and, at the same time, set in motion powerful technology development forces for increased coal utilization. At this point, the price of coal was approximately $1.50 per MM Btu, while that of oil approached $6.00 per MM Btu. This huge price differential offered an incentive for developing new coal technologies including liquefaction, gasification, fluidized bed combustion, coal-water mixture (CWM), etc. The CWM technology was of particular interest to the utility industry, which consumes more than 80% of the coal mined in the US for power generation. In this technology, a powdered coal is mixed with water to produce a stable slurry that can be stored for an extended period of time without settling and is readily pumpable. Various reagents in the amount of approximately 1% by weight of the coal are added to inhibit sedimentation during storage and to reduce the viscosity for atomization in burners. The coal loading is in the range of 50 to 80% solids, depending on the particle size and its distribution. With very fine particles, it is difficult to increase the coal loading above 50%, and a binomial distribution helps increase the coal loading. The presence of water consumes a relatively small amount of the heating value of the coal during combustion; for a coal-water slurry containing 30% water, less than 5% of the calorific value of coal is usually consumed due to evaporation. The major advantage of burning coal as a slurry is that many of the oil-burning furnaces can be converted to coal-burning with a minimum retrofit cost. Considering the substantial price margin between coal and oil, the pay- back period for retrofit costs can be relatively short. Nevertheless, the retrofit cost can be further reduced if low-ash coals are used as feedstocks for CWM; there will be savings in the capital cost of installing bag filters or electrostatic precipitators that are designed to remove the ash from the combustion gases. Furthermore, the use of low-ash coals can minimize derating of boilers because the velocity of the combustion gases can be maximized without creating erosion

Jan 1, 1991

By E. A. Domínguez

La Espingarda kaolin deposit was formed by ?in situ? alteration of volcaniclastic rhyolitic rocks belonging to Jurassic Marifil Formation. Three volcanic lithofacies were identified: porphyric biotitic ignimbrites (PBI), a coarse lithic ignimbrite (ILG), a fluidal intrusive rhyolite (RFI), all kaolinized in different degrees. The alteration covers an ellipsoidal area of ~20,000 m2, with its alteration decreasing downwards and disappearing at 8-12 m from the surface. The length to thickness ratio of alteration is 60:1. An erratic silicification was found as small zones (<3 m2) of quartz stockwork. The mineralogy is kaolinite + quartz + feldspars + Fe-oxyhydroxides ± halloysite ± illite and interstratified I/S, although the relative proportions change in each rock type. Due to coarse grain size and high quartz content, natural kaolin is unsuitable for slip cast ceramics, unless beneficiated. Washed ILG kaolin has the best rheological properties and behavior in casting and drying for its mineralogy, particle size distribution, thin texture and surface activity, as well as a certain fusibility and firing kinetics. PBI and RFI kaolins, due to finer texture, smaller grain size, lower surface activity, and higher halloysite amount, have worse slip viscosity and thixotropy that are due also to soluble salts. For the same reason, they have good casting and drying properties, but their purity brings about higher softening temperatures and water absorption. Several parameters are used to discriminate between hypogene or supergene kaolinite genesis. The hydrothermal genesis is mainly sustained by fluid inclusion temperature (150-180ºC) in quartz veinlets, while weathering is supported by deposit morphology, simple mineralogy, kaolinite stable isotopes (d18O ? 18, 13; dD ? ?59; ?85) and paleoclimatic record. It is concluded that a weathering process formed the main alteration with some participation of steam-heated meteoric waters. Neither the contents of Au, Ag, As, Sb, Hg, nor drill data support the existence of any epithermal gold mineralization at the kaolinite bottom.

Jan 1, 2010

By James A. Madonna

Information gathered while studying geological parameters that govern fine gold deposition in the a1luvial environment can be applied to the technological development and selection of fine gold recovery systems. It is recognized that substantial gold losses occur in the minus 1.5 mn (minus 12 mesh) fraction using conventional sluicing systems. Within the past two decades much attention has been given the design and efficiency of these thick bed concentrators. Studies and comparisons drew information from several foreign and domestic placer operations. These studies concentrated on: 1) the physical and chemical properties of alluvial gold 2) the parameters that control fine gold deposition 3) the geometrical form, components and general working structure of conventional sluicing systems 4) the physical properties which contribute to the overall efficiency of a sluice system 5) the comparison of three major styles of conventional sluicing systems. A dramatic improvement regarding fine gold recovery and water quality was observed when using modern multiple-channel sluicing systems in place of conventional single-channel sluicing systems.

Jan 1, 1988

By Lloyd M. English

During the past 30 years, the research arena has changed significantly in all fields, mining included. The relative representation of the topics of research, how the research is performed and who has performed the research have all changed. Although there are a lot of subjective ideas and anecdotal "analyses" concerning these changes, there has been no actual study of these trends to verify the facts. This study was divided into six five-¬year periods for the purposes of comparison. Some of the trends analyzed are co-authorship, research areas, research sponsors and cooperative efforts. In today&apos;s research arena, 30 years is a long time. Technology has advanced more rapidly than in any time previously and that pace continues to increase. There are more researchers, more tools for research and an increasing number of companies, agencies and institutions engaged in research. The pace of change and its impact is often discussed and debated at professional meetings and in institutions that engage in research. These conversations will often influence the research decisions of the participants in directions to take or subjects to pursue.

Jan 1, 1995

By R. D. Rein

The Henderson mine is a large continuous panel caving operation in central Colorado. The orebody is roughly 760 m by 910 m and will be extracted from two production levels. Each level is divided into three north - south panels 244 m to 317 m wide. The general arrangement of a production level and its associated undercut drifts is shown in Figure 1. The production drifts (3.6 m by 3.7 m) are spaced 24 m apart center to center and have undercut drifts (3.4 m by 3.8 m) 17 m directly above them. The production drifts are interconnected with drawpoint crosscuts (4 m by 4.3 m) driven on 12 m centers. Drawpoint crosscuts receive substantial brow steel and concrete support prior to undercutting. The brow steel is a 15 cm by 46 cm wide flange beam with a 2.5 cm wear plate to box in the &apos;cave&apos; side of the beam. The concrete is a minimum of 0.3 m thick and the batch design results in 27 500 kPa strength. The finished drawpoint opening is 3.7 m by 3.2 m with a choke area of 3.7 m by 2.4 m. (Figure la) Undercutting and bell formation is accomplished by blasting rings of 7.6 cm holes drilled from the undercut drifts. The rings have a burden of 2.0 m and six rings develop one complete bell.

Jan 1, 1985

By J. T. Carlson

Dolomite is an objectionable impurity in phosphate ores, due to its MgO content. Even though dolomite and apatite liberate at a coarse size, the very similar densities of dolomite and apatite makes density separations difficult. However, if operated correctly, jigging can be a very efficient process. This has led to the development of a unique laboratory scale jig that allows for easy alteration of effective jigging properties. Key jigging parameters have been investigated for the separation of dolomite from phosphate ores. It was found that a pulsation rate of approximately 200 pulses/minute worked best for separating dolomite from apatite.

Jan 1, 2010

By J. R. Muhm

Environmental due diligence investigations, also known as Phase I Environmental Audits, are increasingly important to major investors, venture partners, or prospective purchasers or sellers of mineral properties. Knowledge of environmental risks associated with a property will allow better appraisal of a property&apos;s overall desirability prior to completing the transaction. It is possible that the cost of environmental cleanup could exceed the value of the property. Topics appropriately investigated include permit status, water quality, past use of the property, hazardous materials management, storage tanks, public records pertaining to the environmental character of the property, and environmental costs.

Jan 1, 1991

By D. C. McLean

Early studies in the carbonate leaching of uranium ores led to the investigation of pressure leaching as a method to enhance the dissolution of the metal values in this system. The initial investigations conducted at the University of British Columbia supplemented by the work done by the group at the A.E.C. Grand Junction pilot plant, all established the higher efficiency and better economics of the alkaline pressure leaching system. This work was conducted either in laboratory autoclaves using oxygen overpressures, or in pilot plant sized horizontal autoclaves with air as an oxidant. Both of these testing vehicles utilized agitation provided by relatively high speed impellers (300 to 700 rpm), which apparently enhanced the oxidation of the uranium to the hexavalent state. The installation of vertical autoclaves on a production scale, which through design considerations require relatively slow agitator speeds (68 rpm), led to the studies of the utilization efficiency of the air addition and the use of catalytic type oxidants to effect a greater usage of the oxygen added through the aeration system. In addition to investigating the effects of the chemical additives, studies were also conducted on the effect of different impeller configurations on the extraction of uranium, which in itself is a measure of oxidation efficiency. The material presented in this paper was compiled in the Ambrosia Lake district of New Mexico, and the majority of the data was obtained in the Homestake- Sapin Partners mill, located at Grants, New Mexico.

Jan 1, 1960

By G. W. Phelps

This discussion deals with the geology o£ the Tennessee-Kentucky ball clay deposits; their mineralogy and chemistry; methods of finding, mining and refining; and makes observations on reserves of and future prospects for these clays. Ball clays appear in ceramic formulas for manufacture of vitreous china sanitary-ware, electrical porcelains, floor and wall tile, dinnerware and artwares(l). This type of clay is used for refractory (2) and abrasive products(3), and as a suspending agent for ceramic glazes(4), engobes(5) and porcelain enamel slips(6). Although some ball clay is imported, probably 85-90 percent of ball clay used in the USA(7) is obtained from deposits in Weakley, Henry and Carroll Counties in Tennessee(8), and in adjacent Graves County, Kentucky(9). The term "ball", as applied to clay, probably derives from former mining practice in England whereby conveniently-sized blocks of clay were cut from moist clay beds by spading(10), The expression "ball" has no other technological significance. ASTM(11) defines ball clay as "a secondary clay, commonly characterized by the presence of organic matter, high plasticity, high dry strength, long vitrification range, and a light color when fired." This, and similar(12) definitions, applies to English ball clays but is not satisfactory for many American ball clays. Geology of the Tennessee-Kentucky Ball Clay Deposits Figure 1 shows that most commercial clay areas of the eastern United States lie near the edge of the Coastal Plain; a skew-shaped configuration of its landward boundary outlines the Mississippi Embayment(13), Near the apex of this basin a broad plain, with a rolling, Bullied topography(14), dips gently westward from the Tennessee River to bluffs at the edge of the Mississippi River flood plain.

Jan 1, 1972

By V. R. Degner

The trend in both the metallic and non-metallic mining industry is toward processing increasing y higher tonnages of lower grade ores, thereby resulting in a growing interest in progressively larger volume individual machines in the flotation circuit. Because of the relatively high cost of large flotation machine development and evaluation, it is important to establish a rational basis by which the existing smaller flotation machine experience can be extrapolated to establish the configuration of the next larger family of flotation machines with a minimum of development modification in the field. This paper reviews the R&D approach used in translating smaller flotation machine experience to establish the configuration of a 1000 cu ft (2B.3 cu m) induced air machine. Internal cell hydrodynamic principles, which governed this extrapolation, and cell (liquid) surface and froth management considerations will be reviewed, and prototype machine mill tests will be summarized.

Jan 1, 1978

By F. Grajales Muñoz

The Ciénega project, now under construction by Peñoles group, is located in the mountains of northwest Durango, Mexico. The gold and silver deposits are found in a 4-km (2.5mile) long system of veins. An underground operation will be built. The planned 1 kt/ d (1100 stpd) process plant will contain a countercurrent cyanide circuit for oxide ore and a flotation circuit for sulfide ore. Considerable resources have been devoted to the development of infrastructure in the isolated location, as well as for the mitigation of environmental impacts. Startup is scheduled for the third quarter of 1994. Ciénega’s ore bodies consist of quartz veins containing gold, silver, lead and zinc values. The vein thicknesses vary from 2 to 10 m (6.5 to 33 ft) with a northwest trend and nearby vertical dip. Veins are hosted in Tertiary volcanic rocks. The district has intermittently been explored since 1971 by Penoles. After the third series of diamond drilling and direct development work on vein, Peñoles decided to begin project construction.

Jan 1, 1994

By John A. Reeves

The Coal Creek Preparation Plant is the coal cleaning facility of Mid-Continent Coal and Coke Company. Located three miles west of Redstone, Colorado, and at an elevation of 8,040 feet the preparation plant cleans coal produced at the company&apos;s Dutch Creek Mine and L. S. Wood Mine. (See Fig.1.) Current production is at the rate of 2500 tons per day clean coal. Plant capacity is rated 350 TPH raw coal. Reject rates vary with mining conditions from 8% to 25%. Factors which influenced the design of the coal cleaning facility were the characteristics of the coal, the location in a tourist oriented resort area, and the extreme winters of the area. Lack of areas suitable for settling ponds and the necessity of keeping effluent waters out of the fishing streams dictated a closed circuit plant. Severe winter conditions, which last six to seven months out of the year and temperatures which have dropped to 40° F. below zero on occasions demanded that the plant be totally insulated and enclosed indoors.

Jan 1, 1967

By L. C. Stone, J. C. Parr, T. N. Anderson, D. Metcalf

An increase in the importance of low-grade materials as sources of gold and silver has led to the need for improved analytical methods on leach solutions, ores and various effluents, residues, and in-process streams. These are required by plant operators as guides to process control and for accounting pldrposes. Atomic absorption spectrophotometry (AAS) and AAS preceded by pre- concentration have been applied successfully to resolve many such problems. However, AAS is no panacea and imprecision or loss of accuracy can result if adequate precautions are not taken to prevent losses and eliminate interferences. The recent methods are discussed together with methods for overcoming difficulties. Comparisons with more classical analytical techniques are also given.

Jan 1, 1981

By Timothy G. Shuttleworth

In the late1970s, mineral processing plants producing copper sulfide concentrates and oxide leaching feedstocks went online with operating tonnages that exceeded traditional installations. This created a need for larger capacity belt conveyors and apron feeders. Apron feeders are slow moving, heavily loaded conveyors that feed a deep bed of ore from a bulk holding location (a bunker) as a controlled stream, often into a large capacity crusher. Apron feeders use sprockets and caterpillars, while belt feeders use rubber belts with motorized head pulleys to convey the material. The damaging of belts and clogging of transfer points from tramp iron required larger magnetic separators and the removal of tramp metal cannot prevent these serious problems. Corrective action at this location is important for two reasons: Removing tramp iron early in the ore stream protects downstream equipment from potential costly damage. Installing apron feeder magnets before the primary large capacity crusher means that secondary and tertiary crushing steps (made up of multiple parallel units) operate more efficiently, since tramp metal is removed at the onset.

Jan 1, 2011

By Kim A. Lapakko, William W. III White

The ASTM standard method D 5744-96 was developed to provide a well-defined kinetic test protocol, conducted under controlled reaction conditions, with the intent of generating reproducible results. A preliminary indication of the repeatability of a simplified alternative method (MN DNR method) was assessed and results from the MN DNR method were compared to those from the ASTM method. To provide an indication of repeatability, six samples from three different rock types were subjected to dissolution testing for periods of 20 to 68 weeks. Drainage pH values at the end of the tests were within 0.5 units in all cases and within 0.05 units in four cases. The percent difference from the mean for overall sulfate rates from paired cells ranged from 0.04 to 1756, with a median of 6.3%. Nine samples from four rock types were subjected to both MN DNR and ASTM tests for periods of 20 to 132 weeks. The difference in final drainage pH values forsix of the samples was less than 0.1 unit, and these values were less than pH 3.65 or greater than pH 8.0. The final drainage pH difference for the last three samples ranged from 0.15 to 0.4, and these values were ill the pH range of 6.2 to 7.2. The sulfate release rates for the periods of record for the MN DNR method were 70 to 150% of those for the ASTM method, with a median value of 110%. Sulfate and calcium release from the two methods was not substantially different. The MN DNR method typically produced higher magnesium release.

Jan 1, 2000

By Timothy C. Smith, S. Andrew Jacobs, Douglas E. Gonzales, Sandy J. Schexnailder, Carl Urban

At the Schwartzwalder Mine (Cotter Corporation, Golden, Colorado) ion exchange resins have been used in mine water treatment experiments. The experiments, initiated in October 1979 and currently in progress, demonstrate that mine water uranium values can be reduced to levels as low as 10 ppb. Duolite® A-101D, a strongly basic anion exchange resin, was tested in laboratory-scale columns (1 inch diameter) initially and then in pilot-scale columns (4 inch diameter), using both co-current and counter-current operational modes. Observations and conclusions resulting from these studies can be summarized as follows: (1) Duolite® A-101D appears to have successfully extracted uranium at influent levels of approximately 2.0 ppm over multiple loading and elution cycles, consistently producing effluent concentrations as low as 10 ppb uranium. (2) Based upon present test data, counter-current regeneration appears to be the key to successful operation. Upflow loading and downflow elution are the preferred modes for service and regeneration, respectively. (3) Filtration is a necessary pretreatment to remove fine clay and inorganic precipitates that will foul the resin, causing excessive pressure drop and poor loading and elution kinetics. (4) This waste treatment application is unique in that it offers the opportunity to recover some or all of the capital investment in the treatment facility

Jan 1, 1980

By A. O. Drysdale

From a mining point of view the Portland Cement Industry falls logically into the Industrial Minerals Division. Portland Cement manufacture requires four principal materials, lime, silica, alumina and iron. The usual source of lime is limestone. Silica sources include clay, shale, quartz, sandstone and sand. Alumina is usually provided by selecting a silica source with a suitable alumina content such as clay and shale. "Cement Rock" is occasionally found in which all four ingredients are found in natural combination or sufficiently close that only minor adjustments are required. In the absence of iron oxides a white cement can be produced. All raw materials used in Canada contain some iron and in most plants further minor quantities of iron are added to promote fluxing, hence the typical grey colour of Portland cement. The modern period in cement manufacturing started in 1824 when Joseph Aspdin of England patented a process for the manufacture of a modified hydraulic cement which he called "Portland Cement". Hydraulic cements are cements that harden under water. Although natural cements, made by calcining and grinding natural argillaceous limestone were produced in Canada as early as the 1830&apos;s, Portland cement was not produced in Canada until 1890, about 18 years after it was first made in the United States.

Jan 1, 1964

By Richard R. Klimpel

INTRODUCTION The process of froth flotation as a means of upgrading the quality of coal by removing water and/or ash and/or pyrite has been receiving increasing attention since the 1960&apos;s by the world-wide coal industry. Historically, coal preparation practice has involved the use of primarily gravity based separation techniques, screening, and/or water washing in removing the larger fragments of inert material from coarse raw feed coal. The preparation, handling, and cleaning of raw coal finer than 500 micrometers was deliberately avoided. Most often the fine coal that was produced would be judiciously incorporated into the bulk of the larger sized cleaned coal product or simply discarded. Since the 1960&apos;s and accelerating in the 19701s, the desirability of performing some type of coal cleaning or upgrading on the finer raw coal feed materials has been increasing. There are several apparent industrial incentives for implementing fine coal ((500 urn) processing alternatives including the gradually increasing value of the normally lost fine coal product; the additional amounts of fine coal being produced due to changes in mining techniques; the need to appropriately dewater and upgrade fine coal streams in order to meet increasingly stringent environmental regulations on water quality. storage of fines, etc.; the increasing need to lower undesired sulfur and ash contents in some coal end-use applications; and the increasing development of speciality coal use processes that require inherently finer and cleaner coal particles such as in gasification, liquefaction, fuel injection, etc. Looking at the over 70 years of mineral oriented fine particle separation successes at the industrial level, the use of froth flotation in fine coal would also seem to be a natural process to utilize at the industrial level. In the world of strictly mineral processing technology and, more importantly in the world of mineral commodity economics, there are clearly a number of advantages to using froth flotation that are well documented and universally accepted. These include: the relatively low capital and space requirements required when compared to other process options available to upgrade minerals; the extreme flexibility inherent to the froth flotation process both with regard to throughput and differences in raw feed characteristics; the general availability of a wide range of equipment sizes and reliable suppliers of equipment; and the general availability of appropriate and relatively cheap chemicals that are the major driving forces behind mineral flotation. Just how has the implementation of froth flotation fared in coal upgrading over the last 20 years? It is towards answering this question that this paper will be oriented. As such, the discussion will be somewhat philosophical. The comments offered are based on the personal experiences of this author who has worked in the basic side of inventing new flotation chemistry and flotation analysis techniques with the major overriding goal of implementing this new technology successfully at the industrial level in both coal and mineral flotation throughout much of the world. This paper is not intended to be a complete reference to all published scientific work on the subject of industrial froth flotation of coal. As a brief background to what will be expanded upon in this paper, the author would like to offer the following summary comments: 1) the industrial practice of fine coal froth flotation is qualitatively similar but quantitatively different than industrial mineral froth flotation when viewed from either research considerations, operating viewpoints or economic evaluation criteria; 2) it is clearly possible to invent and/or develop new technology (such as new chemistry and/or equipment) that can do statistically better than current plant practice in well-controlled tests in either coal or mineral flotation circuits; 3) there is so much inherent variability to the raw feed coal streams and other operating parameters that even if there is a documented technical incentive to implement some new piece of technology, the people, instrumentation and control methodology required are often simply not available or not economic unless the potential payoff is quite large.

Jan 1, 1988

By L. E. Sausa

Measuring the specific gravity of a pulp in concentrating plants is a relative, y simple procedure, but the desired accuracy is often difficult to attain. Many plants use the system of weighing a known volume of the pulp on a scale. In many cases, this method is quite accurate, but obtaining a true sample presents quite a problem. This is especially true when the pulp tends to froth, the solids settle rapidly, or the flow is fast at the point of sampling. There is also the problem of the operator neglecting to take the sample because of duties elsewhere, or, if everything Is apparently running smoothly, he assumes the pulp density is also all right. In 1952, the Pine Creek plant of Union Carbide Nuclear Company began working on a better method for measuring and controlling pulp density. After an extensive investigation of different methods for measuring the specific gravity of moving pulps, the Research Department decided that the bubble tube method appeared to be the simplest and most promising for our applications. It consists of two bubble tubes, of different lengths, submersed in a pulp. Low pressure air is supplied to, the two tubes through sight bubblers to maintain approximately equal volume of air to each bubble tube. The bubble tubes are also connected to a differential manometer, one to each side. Therefore, knowing the vertical distance between the outlets of the bubble tubes, the manometere can be calibrated to measure directly the specific gravity of the pulp. (Figure No. 1 illustrates this model.)

Jan 1, 1960