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By Wayne Hilgedick

Peabody Coal Company is the nation's largest producer of coal and is headquartered in St. Louis, Missouri. In 1978 Peabody divisionalized its operations to establish more responsive and effective mine management. Peabody Coal Company's Arizona Division operates two large surface mines, the Black Mesa and Kayenta Mines. Both are located in northeastern Arizona, approximately 30 miles from Kayenta, Arizona, on the Navajo and Hopi Indian Reservations. The Black Mesa Mine produces 4 to 5 million tons of coal per year and the Kayenta Mine produces 7 to 8 million tons per year. Coal from the Black Mesa Mine is transported 274 miles by coal slurry pipeline to the Mohave Generating Plant located on the Nevada/Arizona border at Bullhead City, Arizona. This plant is jointly owned by Southern California Edison, the Los Angeles Department of Water and Power, Nevada Power Company and Salt River Project Agriculture Improvement and Power District.

Jan 1, 1983

By S. K. Kawatra, S. J. Ripke

During 1998, 62 million tons of iron ore pellets valued at $1.7 billion were shipped from U.S. mines. These pellets are the primary source for domestic iron and steel production. During pelletization, binders are required to improve the green strength of the pellets. Each year, bentonite binder costs a typical plant from $3 million to $5 million. Reducing the bentonite dosage has the advantages of significant cost savings while producing higher quality, lower silica pellets. The authors have recently identified a new and important bentonite binding mechanism, the fiber theory. During this study, procedures were developed that took advantage of the ability of bentonite to form fibers, and resulted in a dramatic decrease in the necessary bentonite dosage. Acceptable quality pellets were produced with the bentonite dosages reduced in half from 14 lb/lt (0.66%) to 7 lb/lt (0.33%).

Jan 1, 2001

By A. Goulsbral, D. Harrison, R. Dunne

A programme to reduce cyanide consumption at the New Celebration Gold Mine, situated 32kms South of Kalgoorlie, Western Australia, commenced in the latter part of 1990. Cyanide speciation studies at the time identified that the major loss and consequently consumption was due to the hydrolysis of cyanide, The use of artesian hypersaline waters of high magne­sium content on gold plants in Western Australian Goldfields has meant that economically viable operating pH values during cyanide leaching are around 8.9 to 9.5. The high ionic strength in these waters does to some degree retard the hydrolysis reac­tion. Notwithstanding this, the formation of hydrogen cyanide and its loss due to volatilisation is the cause of the major loss of cyanide on most gold plants in the Kalgoorlie Goldfields. This paper discusses the aspect of oxygen addition during cyanide leaching to decrease cyanide consumption and increase gold recovery.

Jan 1, 1996

By Rod Knopp

Each year approximately 5,000 tons of a sulfide concentrate grading 1 ounce gold per ton are produced at this operation. The material is refractory and as shipping costs and custom smelter charges increase, options to process it on site are becoming more attractive. Three methods of treating the ore have been investigated in detail. The first option of ultra fine grinding the concentrate produces a 35% gold recovery with relatively low capital and operating costs. A second option of pressure oxidation is very capital intensive and has high operating costs but recoveries in the order of 95% are achievable. The final option, biological leaching offers both capital and operating costs which are less than pressure oxidation, while still achieving 90% gold recoveries. Bioleaching of gold concentrates has been gaining acceptance but there is a risk factor associated with this or any emerging technology.

Jan 1, 1990

By Harry M. Castrantas

Caro's acid (peroxymonosulfuric acid-H2SO5) is used to destroy cyanides in tailings slurries due to its rapid reaction and low capital and operating costs. Due to its instability, Caro's acid is generated at the mine site using a hot or cold process. The majority of gold mines that use Caro's acid use the FMC hot process. In the FMC hot process, it is not necessary to cool the Caro's acid to minimize decomposition since it is used immediately upon generation. In contrast, the cold process uses a chiller to cool the raw materials and product. This paper will show, using laboratory and field data, that there is virtually no difference in Caro's acid yield between the cold and the FMC hot processes. Caro's acid yield depends on the mole ratio and concentration of sulfuric acid and hydrogen peroxide and is independent of operating temperature.

Jan 1, 1997

By Dave Willis

Two hundred and thirty million tons of coal! That is what the Burlington Northern (BN) railroad anticipates that it and the Chicago North Western railroad (CNW) will haul out of the southern Powder River Basin l (see map on page 966) in 1994. Put that 209 Mt (230 million st) into railroad perspective. Current average train size is 113 cars. With each car transporting 94 t (103.5 st) of coal, each train averages 10.6 kt (11,700 st) of coal. Simple division shows that 209 Mt (230 million st) of coal equals 19,658 trains a year. So BN and the CNW must move 54 loads a day out of the Basin. Every load, of course, has an empty side of its cycle. This means that 54 loads a day translates into 108 total trains a day, or 39,420 trains a year through the southern PRB.

Jan 1, 1994

By B. Das, B. K. Mohapatra, K. S. Narasimhan, S. K. Bhaumik

The iron ore slimes (d80 ~ 50 µm) generated at Noamundi iron ore washing plant (TISCO), assaying 57% Fe, 4% Si02 and 8.3% A1203, have been subjected to hydrocyclone studies. The effects of different design and operating parameters have been studied to get a product suitable for subsequent use in the sinter mix. The studies indicate that it is possible to obtain a product containing 64% Fe, 1.4% Si02 and 3.5% A1203 with a recovery of 49% iron. The process involves classification followed by separation in a hydrocyclone. By this method, the overall yield of iron ore production in the plant is expected to increase by 8%. Microscopic examinations and X-ray diffraction studies were conducted to elucidate the mineralogy and liberation characteristics of Noamundi iron ore slimes. This paper deals with the findings of the investigations.

Jan 1, 1993

By Robert C. White, Kurt W. Keskimaki

The Henderson Mine is a large underground panel cave molybdenum mine located near Empire, Colorado. Diesel equipment is employed for all development work, as well as production ore handling. Due to fluctuating demand for molybdenum, production rates have varied from 16,000 to 32,000 metric tons per day (18,000 to 36,000 short tons per day) over the last four years. In order to obtain the maximum air flow for minimum power cost the Henderson mine has completed the following modifications to the ventilation system: removal of large intake fans, elimination of underground booster fans, shutdown of tunnel exhaust fans, completion of the number 5 exhaust ventilation shaft, and use of partially debladed main exhaust fans. These changes have reduced the annual cost of mine ventilation at the Henderson Mine by over 80 per cent, a very critical cost savings in today's competitive molybdenum marketplace.

Jan 1, 1989

By Cherie M. Tilley

The Stillwater Mining Company has acquired a Tunnel Boring Machine to expedite the mine development operations at its Stillwater Mine. The machine purchased by Stillwater was previously owned by Melbourne Metropolitan Board of Works, Melbourne, Australia. The machine was manufactured by Robbins Co. of Kent, Washington. The Stillwater Mine has an extensive system of footwall laterals which are required by the mine development scheme. The initial mine plan called for in excess of 30,500 linear meters (100,000 linear feet) of footwall lateral drifts at a cross section of approximately 3.0 meters in width by 3.75 meters in height (10 feet x 12 feet). A Tunnel Boring Machine was selected for this work. The Robbins machine number is 146-193. This machine is approximately 4 meters (13 feet 5 inches) in diameter. The machine was modified extensively prior to commencing operation at the Stillwater Mine.

Jan 1, 1989

By G. G. O. O. Uwadiale, M. A. U. Nwoke

The flotation of Azara barite tailings was investigated using crude rubber nut oil (RNOc), refined rubber nut oil (RNOR) and oleic acid as fatty acid collectors. Chemical characterization of the rubber nut oils indicated high acidity, high saponification values and high iodine indexes, which are indicative of a high fatty acid compositions. The Azara barite tailings analyzed 64.6% BaSO4 by atomic absorption spectrophotometry. Gravity and magnetic separation techniques yielded a BaSO4 concentrate with a grade of 74.8% and a recovery of 92.5%. Barium sulfate grades of 90.0%, 92.8% and 94.8% with recoveries of 80.3%, 85.4% and 86.7% were achieved with collector dosages of 750g/t crude rubber nut oils, refined rubber nut oils and oleic acid, respectively. Hence, rubber nut oils are effective fatty acid collectors for the flotation of barite tailings.

Jan 1, 1998

By H. L. O. Huyck

A microcomputer equipped with spreadsheet, data management, and FOR- TRAN programming software is used to update and quantify a model for ranking seventeen economically important metal commodities in terms of market potential (Huyck, 1982). Weighted factors are used for the ranking: (1) demand growth trends, (2) end-use substitution vulnerability, (3) recycling trends, (4) price trends, and (5) cumulative demand versus world reserves factor. Price trends are defined using (1) price-vs.-year slopes (with prices in 1967 constant dollars) and (2) standard deviations of slopes (to indicate price volatility). Substitution vulnerability is defined as the percent of end-use which is replace- able by other commodities, and a factor indicating the likelihood of such substitution. Factors are normalized and linearly combined to produce a ranking of the seventeen metals. Microcomputer software is essential for organizing data and calculating statistical parameters and linear combinations of factors.

Jan 1, 1986

By A. Pratt, G. Deschênes

This paper quantifies the detrimental effects of pyrite and stibnite on gold leaching kinetics. Means to reduce the effect of these minerals by addition of lead nitrate and oxygen are presented. A significant retarding effect was observed when the pyrite concentration was 20%. A pretreatment with lead nitrate reduced the extent of gold surface coating by Ag, Fe and S. Stibnite, at a concentration as low as 20 ppm, significantly affected the gold leaching kinetics through forming a passivating film of Sb(V) oxide possibly Sb2O5. The most practical way to alleviate the effect of stibnite is to leach the materials at a moderate pH with addition of lead nitrate. A low temperature reduced the efficiency of lead nitrate to counteract the effect of stibnite. The increase of dissolved oxygen is not favourable to alleviate the coating of stibnite on gold.

Jan 1, 2004

By Isabel F. Barton, MarkD. Barton, Christian A. Rathkopf

Molybdenite is the world’s principal source of rhenium (Re), a critical element in multiple high-tech applications. However, the Re contents in molybdenite vary by orders of magnitude on scales ranging from single grains to whole deposits. In order to better understand the systematics of this variation and what geochemical factors control molybdenite Re concentration, and hence overall Re resources, we examine global patterns in molybdenite Re contents through a compilation of > 3000 measurements of Re in molybdenite from > 700 mainly ore-bearing moderate- to high-temperature hydrothermal systems of different types. Our results are similar to but expand on those of earlier studies. Rhenium concentration in molybdenite has a lognormal distribution and varies systematically with type of geologic system, intrusive lithology, and Mo grade. The lowest-Re molybdenite occurs in greisens (geometric mean 1 ppm ± a multiplicative standard deviation of 9), quartz vein-hosted W-Sn deposits (2 ± 5 ppm), unmineralized granites and granodiorites (12 ± 8 ppm), intrusion-related deposits (24 ± 8 ppm), and porphyry W-Sn deposits (16 ± 11 ppm). Rhenium is most enriched in molybdenites from volcanic sublimates (23,800 ± 5 ppm), with skarn Fe and Au (560 ± 5 ppm and 540 ± 3 ppm respectively) and porphyry Cu and Cu-Au deposits next (470 ± 4 and 430 ± 7 ppm respectively). Among porphyries, skarns, and quartz vein-hosted deposits, Re is most highly concentrated in molybdenites from Cu and Au systems and its concentration decreases systematically through Cu-Mo, Mo, Sn, and W deposits. In nearly all cases, molybdenites from systems associated with intermediate igneous rocks contain more Re than molybdenites from systems of the same type with more felsic rock associations. The disparity between Re contents of molybdenite in felsic and intermediate systems is largest for porphyries, quartz vein-hosted, and skarn deposits and is near zero for subeconomic or barren granite and granodiorite Mo systems; felsic intrusion-related deposits have slightly higher molybdenite Re than their equivalents associated with intermediate intrusions. In most systems, molybdenite Re content does not correlate with metal grade, but may have an inverse correlation with Au grade in intrusion-related deposits (based on a small number of data points) and does exhibit a strong inverse correlation with deposit Mo grade. Dilution of Re through larger amounts (higher deposit grades) of molybdenite explains about 40% of this correlation, but the relative enrichment of Re in molybdenite from low-Mo deposits must also reflect some selective enrichment of Re/Mo in porphyry Cu systems compared to porphyry Mo systems. We found no evidence for secular increase or other systematic temporal variation in molybdenite Re content. The data regarding the use of molybdenite Re content as a proxy for mantle influence are ambiguous. Nearly all observed empirical correlations can be traced back to differences in redox state and sulfide concentration, the two geochemical factors identified here and by previous experimental work as the controlling influences on Re mobility under hydrothermal conditions. Hydrothermal systems with reducing conditions (W- and Sn-rich) tend to have low molybdenite Re even though compiled whole-rock data indicate that their source rocks have as much or more Re as those of more oxidized systems (e.g., Cu-rich). Vapor-phase exsolution, crustal assimilation, and mixing with external fluids may all enrich molybdenite Re concentrations in individual deposits and deposit types, but their extent and importance in overall hydrothermal concentration of Re is uncertain. Thus, it appears that the available molybdenite Re resource in an ore deposit largely depends on how the deposit’s redox and sulfidation conditions have varied over time and space during the timespan of hydrothermal activity. Oxidized, high-sulfide conditions tend to concentrate Re in molybdenite, whereas reducing conditions tend to leave Re dispersed at low concentrations in the bulk rock.

By M. Fig, J. Brune, C. Strebinger, G. Bogin, J. Grubb

"Although it is known that explosive gas zones can form and sustain combustion in a longwall coal mine gob, a complete account of the interaction of the methane flame and gob remains obscure. The goal of this study is to examine the impact of several gob parameters such as packing orientation and void spacing. The impact of ignition location is also detailed. A series of experiments with methane flames were performed in a horizontal cylindrical quartz reactor. The reactor was tested both empty and packed with an obstacle wall consisting of spherical glass beads to investigate the impact of various gob parameters. Results demonstrate that methane burning velocity is sensitive to all parameters investigated, and most sensitive to packing density and ignition location. INTRODUCTION The explosion at the Upper Big Branch Mine in 2010 and several other underground longwall coal mine fires and explosions show that explosive gases can mix with fresh ventilation air creating explosive gas zones (EGZs). Investigative reports of these fires have found that EGZs are able to move from the gob and into the active longwall face, endangering workers and equipment [1]. Currently, researchers at the Colorado School of Mines are using computation fluid dynamics (CFD) modeling to predict the location and movement of EGZs in the gob and along the longwall face [2,3]. It is the goal of this work to validate combustion models [4] that include interactions of the flame and the gob for future incorporation into ongoing CFD modeling of EGZs [2,3]. Throughout the years, there has been significant research investigating how flames interact with obstacles and propagate through porous media [4,5,7-11]. Since there is limited access to the gob in a longwall coal mine (Figure 1), the orientation and size of rocks and void spaces within the gob is unknown. Experiments of a simulated gob have been performed in a large scale flame tube in order to investigate the effect of scale (Figure 2) [4,5]. However, the variation in rock spacing, rock orientation, and void location makes it difficult to determine the impact of each of these parameters on methane flame dynamics. This study aims at gaining a stronger fundamental understanding of gob parameters such as spacing, orientation, and ignition location on methane flame propagation. The results from this study will provide a basis for experimentation in larger explosion vessels and will be used to validate methane combustion models for incorporation into longwall ventilation models of EGZ movement [2,3,4]. BACKGROUND Researchers have been studying flame propagation across obstacles and porous media for many years using various fuels and experimental setups [4,5,7-11]. Experiments have been performed in horizontal chambers with both ends open [7], horizontal chambers that are semi-open [4,11], horizontal chambers with both ends closed [8], and vertical arrangements [9,10]. Due to the variation in experimental setups and methods of characterizing flame dynamics such as flame speed, peak overpressure, or burning velocity it becomes difficult to directly compare results. Flame speed is a property of the mixture and refers to the speed of the unstretched flame front normal to the direction of propagation. This study characterizes the flame by its burning velocity, which is the difference between the flame speed and the speed of the unburned gases just ahead of the flame front. However, some references use the terms flame speed and burning velocity interchangeably. In this study, a semi-open quartz cylindrical reactor was used to determine methane burning velocity, which was used as a basis for comparison to determine the impact of obstacle configuration on methane flame dynamics. The quartz reactor also has the unique ability to allow for flame visualization. These features allow for faster combustion model validation [4]."

Jan 1, 2017

By Erich U. Petersen, W. Dan Hausel, William J. Tafuri, Donald M. Hausen, Douglas N. Halbe

The geology of Wyoming has been compared to some of the richest precious metal producing regions of the world including South Africa, Western Australia, and the Superior Province of Canada. These areas are all underlain by ancient cratons that contain some of the oldest rocks on the surface of the earth. But unlike the other cratons, the Wyoming craton (also termed the Wyoming Province) has been greatly modified by Laramide tectonics which thrust slices of the ancient craton through younger Phanerozoic sedimentary rocks. Unfortunately, the nature of the Laramide tectonics left vast regions of the craton at the bottom of the Tertiary basins unavailable to direct examination, but the exposed slices in the mountain cores contain some of the best exposures of Precambrian rock in the world. The cores of these uplifts contain extensive regions of relatively unmineralized gneiss and granite that include scattered fragmented supracrustal belts formed of metamorphosed volcanic, plutonic, and sedimentary rock. Some of these supracrustal belts exhibit similarities to the gold- and nickel- rich terranes of Western Australia. Yet these areas remain incompletely explored. Archean (> 2.5 Ga) gold deposits within the Wyoming craton are primarily confined to greenstone belts and related terranes. These supracrustal belts include relatively narrow vein and shear zone deposits with strike lengths of tens of feet to more than a mile. Some exceptional deposits possess widths of greater than 15 feet to possibly as much as 300 feet, but the majority have widths of± 5 feet. These belts include Cu, Ag, W, Fe, and Cr mineralization, asbestos and pegmatites, and anomalous Ni and Sri in addition to gold. The Proterozoic terrane (<2.5 Ga) exposed in the cores of mountain ranges in southeastern Wyoming includes a thick miogeosynclinal metasedimentary succession which unconformably lies on the Archean craton. These metasedimentary rocks are separated from a predominantly metavolcanic terrane to the south by a major Precambrian suture, or shear zone. The metasedimentary rocks north of the shear include quartzite-hosted stratabound Cu-Ag-Au deposits, Au-Ag veins, and Witwatersrand-type metaconglomerates with isolated gold anomalies. The suture includes scattered base and precious metal deposits in shear zone cataclastics. South of the shear zone, the metavolcanic terrane includes scattered volcanogenic Zn-Cu-Ag massive sulfide deposits, at least one Cu-Au porphyry, and two large layered mafic complexes. The northern edge of one of these layered intrusives yielded some platinum and palladium to miners at the beginning of the 20th century. Phanerozoic sediments host many precious metal anomalies. Some Au-REE anomalies occur in Cambrian conglomerates. Ag-Cu-Zn stratabound deposits are found in bleached Jurassic redbeds in the Overthrust Belt of western Wyoming. The broad Wyoming basins include many enigmatic gold occurrences and anomalies that can only partially be explained by detrital transport in fluvial systems. Many of these anomalies were probably geochemically transported and precipitated. Rocks of Late Cretaceous age along the flanks of the uplifts have numerous black sand deposits. These paleobeach placers contain concentrations of detrital heavy minerals with anomalous titanium, rare earths, and in some cases gold. Gigantic base and precious metal deposits are associated with scattered composite intrusives in the Absaroka Mountains in northwestern Wyoming. This deeply incised volcanic plateau includes several Tertiary age copper porphyry deposits with significant Cu, Mo, Ti, Pb, Zn, Au, and Ag resources that rival the porphyry districts of the Basin and Range Province. In the Black Hills of northeastern Wyoming, disseminated Au-Th-REE mineralization is found in association with fenitized alkalic igneous rock. In addition to disseminated mineralization-scattered gold, fluorite, rare earth and copper veins; lead, zinc, and silver replacement lodes; and tin pegmatites are found in this region.

Jan 1, 1990

By Ryne Goodnough

INTRODUCTION The Wyodak Mine is located in the northern Powder River Basin of Wyoming and is thought to be the oldest continuously operating coal mine in the United States, having recorded yearly production since 1923. The mine has expanded over its 80-plus-year history from an underground room and pillar operation to a modem surface snip mine. Currently, the mine feeds four on-site power plants using conveyors, has a filly automated truck load-out, and ships coal by rail to nearby Wyoming power plants. Annual production of 4.8 MTFT is accomplished on a single-shift basis with 53 fill-time production, maintenance and administrative personnel. In 1990 a decision was made to modernize the coal feed system to the power plant complex with modular conveying equipment. The project involved constructing coal storage silos, overland conveyors, feeder breakers and modular conveyors to follow the mining benches. Total project cost in excess of $25 million was used to construe or commission: A 2,150-TPH 48-inch overland conveyor with horizontal curve A 2,150-TPH 48-inch plant feed conveyor 13 shuttle conveyors, each 150 feet in length 2 front-end loaders equipped with large coal buckets 2 crawler-mounted feeder breakers with compatible discharge conveyor 1 high-lift shuttle conveyor with adjustable discharge height 1 high-angle conveyor 2 4,500-ton coal silos with automated truck load- out capabilities. Conceptually, the plan was to develop 33 MTONS of reserve adjacent to the plant complex and use front-end loaders to tram ROM coal to in-pit feeder breakers. The primary crushing would reduce coal to 6-inch minus and feed through a series of shuttle conveyors to an overland supply conveyor. The overland conveyor would feed the secondary crusher, and the crushed product would be moved to any of four storage silos using conveyors. The in-pit crushing and conveying system is a unique fit for the power plants and provided many interesting challenges for the coal mine to overcome. System integration, maintenance windows, equipment availability and the like had a major role in engineering the system to a power plant complex having no emergency coal stock- pile and limited plant storage. In crediting the designers and operators, the in-pit conveying system at the Wyodak Mine has delivered coal to the power generators since 1994 with little or no downtime. ENGINEERING CONSIDERATIONS Yearly production for the mine is based on 3.5 MTPY which requires about 1,550 TPH on an 8 hour, one-shift- per-day, 5-days-per-week basis. Considering efficiencies for starting and stopping, crew transport, loader trams, and operational delays, a nominal conveyor feed rate of 2,150 TPH or 72% system utilization was used in engineering design application. This ton-per-hour basis is used to size and commission the remaining components of the in-pit system. Coal quality requirements also played a role in equipment selection. The mine reserve grades vertically and coal from Upper and Lower benches needed to be mixed in any given proportion. This required equipment to be filly variable and capable of producing all or part of the 2,150-ton-per-hour requirement. Bisecting the 80-foot- high coal face for quality required construction of a high lift module capable of reaching a single overland conveyor 40 to 50 feet above the pit floor. Designers built a high-lift, adjustable and portable module specific for this application Having two bench systems capable of each producing the required feed for the power plants required computerized control processes to control on-line production. RSLinx, RSLogix, and RSView were chosen as the communication and control software, and a Data Highway Plus serves as the communication backbone. Each of the two in-pit benches are controlled with a data center equipped with Allen Bradley 5-40 processors. This

Jan 1, 2004

By James M. Williams, Michael E. Zellars, Arthur J. Roth

INTRODUCTION Our present estimates of world P205 consumption for all uses show an average annual growth rate of approximately 1.9% per year. The estimated 1985 consumption of 41.5 million tonnes of P205 is predicted to increase to 57.1 million tonnes of P205 by the year 2005, which represents a growth from 157.7 million tonnes in 1985 to 217 million tonnes of phosphate rock in 2005. The present situation of oversupply and extremely weak phosphate product prices is, however, expected to continue in the short term and a normal supply/demand balance may not be achieved until after 1990. The growth in P205 demand, in conjunction with the higher cost of replacing low cost mines as they are depleted, will place economic pressure on world farm economies through the increased cost of phosphate fertilizers. The availability of rock supply is not the challenge; rather, it is the economics of new rock supply. This economic challenge of providing new rock supply is by no means limited to the United States. Even in centrally controlled economies, the cost of rock from domestic mines of marginal quality and economic viability, coupled with burdensome infra¬structure requirements, may result in tonnage additions at very high cost. Although government subsidies to offset higher production cost is a factor in some countries, world production in the next 20 years will be directed, if not governed, by economic factors. Established infrastructure, good quality reserves, and existing operations will provide a competitive advantage for new prod¬uction. It is our opinion that the current world supply structure will remain relatively stable through 2005, in that most of the traditional suppliers will retain supply roles. U.S. rock production, although it is not expected to grow beyond the present level, will still be a major contributor to world supply. Although dramatic changes within mining, fertilizer production and farm economics may appear disruptive, the current supply scenario will, in general, continue with the major issue being production economics. WORLD DEMAND Demand for P205 by region in the world has been estimated using projections of population, crops and yield, arable land, economic and other data as a basis for future demand projec¬tions by Blue-Johnson and Associates, our col¬laborators on a recent multi-client study. Non-fertilizer P205 demands were estimated in the same general manner. These estimates of future demand are based on an assumption of reasonable world political and economic stabil¬ity, debt repayment requirements maintained at workable levels, and resolution of problems in the agricultural economy to a point where normal supply/demand forces can interact freely to provide a basis for continued P205 supply. The results of these studies indicated an average growth rate of 1.9% per year through the year 2005. This growth could be as low as 1.5% per year, or as high as 2.5% per year. It is our opinion that the estimated rate of 1.9% per year is most probable. This moderate growth rate considers both the major economic constraints facing the world agricultural industry today and the general problem of affordability of fertilizers and other crop inputs. WORLD SUPPLY There are adequately defined world reserves of phosphate rock today to provide supply well into the future, even at much higher levels of demand than presently projected. Using the U.S. Bureau of Mines estimates, which are based on stringent criteria for definition of reserves, the present world reserves base is approximately 35 billion tonnes. Identified resources raise the potential for production by orders of magnitude. Potential resources based on geological trends, and including offshore deposits, would again result in orders of magnitude increases in rock resources. Morocco leads the world in reserves base with approxi¬mately 21 billion tonnes. The U.S. is second, followed by the Republic of South Africa and Russia. These four countries control about 87%

Jan 1, 1986

By Tim O’Neil

Until about three years ago, Asarco&apos;s copper business consisted predomi¬nantly of custom smelting of ores and concentrates produced by other mining companies. Since then, the company has been transforming itself into a fully integrated producer in copper mining, smelting, and refining. In doing this, Asarco has lowered its costs and restructured its operations and finances. Now, Asarco hopes to complete this process by spending $260 million over the next three years, to expand and modernize its copper facilities and boost production by some 40%. The capital spending program includes: • $130 million at the Ray mine about 113 km (70 miles) north of Tucson, AZ, to expand mining capacity and install an in-pit ore crusher, mill, and concentrator; • $100 million, to expand the mining capacity at the Mission Complex located south of Tucson, AZ, and refurbish the adjacent, idle Pima mill and concentrator. (Asarco had earlier exercised an option to purchase the Pima mill for about $6 million); and • $30 million, to modernize the copper smelter in El Paso, TX, with a new flash smelting process. The new program is in addition to a recently completed first-phase expansion of the Mission mine, mill, and concentrator. This $13 million debottlenecking of existing operations in¬creased production capacity by 46% or 24.5 kt/a (27,000 stpy), to 79 kt/a (87,000 stpy) of copper in concentrates. In addition, a previously an¬nounced $12 million expansion program at the Ray mine is scheduled for completion in early 1990. It will expand mill capacity, to offset the anticipated effects of increasing ore hardness as the pit deepens. Ray produces 68 kt/a (75,000 stpy) of copper in concen¬trates and an additional 36 kt/a (40,000 stpy) of electrowon copper. By 1992, when Asarco&apos;s expanded and modernized copper facilities are operating at capacity, the company&apos;s mine output will have increased by 67 kt/a (74,000 stpy), to 263 kt/a (290,000 stpy) of contained copper. That will be enough to provide all of the feed required for Asarco&apos;s two copper smelters - by then, both of them modern, state-of-the-art facilities. Asarco&apos;s expansion and modernization program will further reduce costs and provide added assurance that the company&apos;s copper business will be profitable at the bottom of the cycle, according to Chairman Richard de J. Osborne The Ray mine portion of the new program will include construction at the mine of a new mill and concentrator, with a capacity of 18 kt/d (20,000 stpd) of ore. These new semiautogenous grinding mills and large capacity flotation cells will augment the present 27.2 kt/d (30,000 stpd) concentrator located 29 km (18 miles) away, in Hayden, AZ. A 29 km (18 mile) pipeline will be built to carry tailings in slurry form from the new Ray mill to the present tailings pond. Concentrates from the new facility will be shipped by rail to Asarco&apos;s Hayden copper smelter. Ray&apos;s present 27.2 kt/d (30,000 stpd) crusher, adjacent to the open-pit, will be replaced by a 54.4 kt/d (60,000 stpd), portable in-pit crusher and conveying system. This will reduce the more expensive ore haulage by truck. The Ray project is scheduled for completion by 1992. It will increase the mine&apos;s annual output of copper in concentrates by an additional 33.5 kt/a (37,000 st), to 102 kt (112,000 st). And the project could mean an additional 400 jobs at the 730 employee Ray operation. Work at the Mission Complex involves reactivation of the Pima mill and concentrator and expansion of mine output sufficient to provide ore to both the Pima mill and the present Mission mill. In 1985, Asarco purchased the Pima mine, which occupies one end of the Mission pit. The work at Mission will increase its annual capacity by 33.5 kt (37,000 st), to 112 kt/a (124,000 st) of contained

Jan 1, 1989

By William R. Henkle

High-calcium limestone has been intermittently produced from the Monarch quarry since the late 1800s. Initial production was used in sugar processing and for a smelter located near Salida, Colorado. During the 1920s, Colorado Fuel and Iron Company (CF&I) began to exploit the deposit as a source of flux stone for its blast furnace steel operations at Pueblo. The deposit was operated for many years as a captive quarry by CF&I. In 1984, CF&I closed the quarrying operations at Monarch; flux stone was no longer required since the blast furnaces were taken off line in 198 1. In 1994, Colorado Lime Company, a subsidiary of United States Lime and Minerals, Inc. acquired the property. Limestone was mined from the Castle Butte Member of the Leadville Limestone (Lower Missippian). Some high- magnesium dolomite was also mined during WWII from the Redcliff Dolomite Member which directly underlies the Castle Butte. Past production was from both an upper and lower quarry at elevations ranging from 10,000 to 11,000 feet (3,048 m-3,353 m) mean sea level (m.s.1.). The reserves at Monarch quarry are in a dip slope deposit, occupying the northwest slope of Monarch Hill. The structural geology of the deposit is complex on both a region- al and mine scale. The reserves are located on the eastern limb of the Syncline Hill syncline. The beds on this limb of the structure strike to the northeast and dip to the northwest from 30 to 45 degrees. A major reverse fault separates the reserves into an upper and lower deposit. Numerous smaller normal and reverse faults cut the deposit, requiring a flexible mine plan. At the Monarch quarry, the Castle Butte Member con- sists of two stratigraphic units. The Upper unit (Mlu) is a thick-bedded, fine-grained crystalline dolomitic limestone from 0 to 300 feet (0-90 m) thick. The unit generally has an Si02 content of less than 4 percent. A bedded chert horizon and a karst horizon at the base of the unit locally raise the Si02 content as high as 20 percent. The underlying Quarry Stone unit (Mlq) is a light-gray, massive, bedded, calcilutite/micrite limestone from 90 to 130 feet (27-40 m) thick. The Quarry Stone unit is remarkably uniform chemically throughout it&apos;s entire thickness. Average analysis is CaC03 - 95 percent, MgC03 - 3.5 percent, and Si02 - 0.5 percent. The Footwall dolomite unit (Mld) of the Redcliff Member of the Leadville Limestone is a dark gray to black, blocky, medium-to-thick-bedded, calcilutite/micrite dolomite from 150 to 180 feet thick (46-55 m). The upper 100 feet (30 m) of the unit is a low-silica facies with an average Si02 content of less than 0.5 percent. MgC03 content is con- sistent at 39.8 percent throughout the upper 50 feet (15 m) of the unit but varies from 15 percent to 39.8 percent in the lower portion (50 to 100 feet [15-30 m]) of the low-silica facies. The lowest portion of the unit, 100 feet (30 m) to base of unit is high in silica and is not considered chemical grade. Current in-place, surface-minable reserves at the quarry are estimated to be 14 million short tons (st) (13 million metric tons [mt]) of limestone and 50 million st (45 million mt) of dolomite. A 1.4 million st (1.3 million mt) stockpile of finely crushed limestone is also present on the property. Colorado Lime is presently processing approximately 2,000 st (1,800 mt) per month from the crushed limestone stockpile. This material is pulverized and sold to underground coal mines for use as rock dust, some of this product is also used for environmental purposes. Limestone mining is scheduled to resume in spring of 2000.

Jan 1, 2001

By M. Churchill, V. R. Degner

Over the past two decades, declining ore body grades and increased operating and construction costs have provided an incentive for the development of progressively larger flotation machines. As a result, the 1.7-2.8 m3 (60-100 cu ft) flotation machine of the 1950&apos;s and 1960&apos;s is being rapidly replaced by large volume machines. The result has been improved flotation machine specific capacity (tph processed per sq ft of floor space); specific power (hp per tph processed); and process control simplicity. The recent development and satisfactory plant evaluation of the 28 m3 (1000 cu ft) flotation machine allows economic and operating advantages in high tonnage mill applications. Some principles of mechanically induced air flotation and the development of the large 28.3 m3 (1000 cu ft) flotation machine are reviewed here. Performance evaluation in beneficiating metalliferous ores-copper, zinc, and iron-is also given and recently completed plant results provided. Mining lower grade ore bodies results in the need to process increased tonnage in the concentrator. This greater tonnage requires more individual flotation machines, unless individual machine capacity, or cell volume, is increased. As the ore processing rate of the concentrator increases, the benefits of progressively larger individual cells in the flotation section become more significant. The accompanying figure illustrates the advantages of a larger flotation machine. In the figure, the number of machines in the rougher-scavenger section is related to the plant tonnage for three sizes of flotation machine: 8.5 m3 (300 cu ft), 14.2 m3 (500 cu ft), 28.3 m3 (1000 cu ft). The 28.3 m3 (1000 cu ft) flotation machine results in 95 fewer machines, compared to the 8.5 m3 (300 cu ft) machine at 45.4 kt/d (50,000 stpd) feed. The flotation cell reduction is, of course, doubled at a plant production rate of 90.7 kt/d (100,000 stpd). Using fewer flotation machines saves floor space. The 14.2 m3 (500 cu ft) and the 28.3 m3 (1000 cu ft) machines result in a floor area reduction, relative to the 8.5 (300 cu ft) machine requirements, of 134.7 and 418.1 m2 (1450 and 4500 sq ft), respectively, for the 45.4 kt/d (50,000 stpd) rate and 269.4 and 836.1 m2 (2900 and 9000 sq ft), respectively, for the 90.7 kt/d (100,000 stpd) rate. In addition, flotation practice has shown that the "specific power intensity"-power per unit cell volume-is less for increased cell size. This results in a power draw reduction, for the rougher-scavenger circuit, of between 552 and 686 kW (740 and 920 hp) for the 14.2 and 28.3 m3 (500 and 1000 cu ft) machines, respectively, when processing 90.7 kt/d (100,000 stpd). Fewer cells improve machine control and simplify maintenance needs while reducing floor space and power requirements. These are tangible incentives for the development of large flotation machines for high tonnage applications. Flotation Principles The accompanying figure shows how key hydrodynamic regions influence the performance of a mechanically induced air flotation machine. Air is induced into the fluid from the top of the cell, due to the mechanical action of a rotating impeller. The impeller circulates the liquid, or pulp, from the bottom of the cell. The result of these dual flow paths is the mixing of the two fluids and one solid phase in the highly agitated region where physical contact between the air bubbles, and the solid particles to be floated, is accomplished. The particle to be floated is assumed to have a properly prepared hydrophobic surface prior to flotation. Upon leaving the three-phase mixing region, the "floatable particle-air bubble matrix" and

Jan 7, 1982