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By Robert L. Cheek, Sid McIlveen

Sodium sulfate is an important industrial chemical. As recently as ten years ago it was produced and consumed in the United States in quantities exceeding 1 Mtpy. Since then, both its production and use have declined; however, approximately half the production still comes from natural sources. [Fig. 1] illustrates the history of production of natural sodium sulfate in the United States. Production of natural sodium sulfate from various types of deposits is the main source of this chemical in Canada and Mexico, and probably in Argentina, Chile, Iran, Spain, and the Russian Republics. MINERALOGY AND PHYSICAL PROPERTIES Sodium sulfate is widespread in occurrence and is a common constituent of many mineral waters, as well as seawater. Atmospheric precipitation contains sulfate; it is one of the major dissolved constituents of rain and snow (Davis and Dewiest, 1966). Many of the saline lakes throughout the world contain varying amounts of sodium sulfate. Because sodium is usually the dominant cation, some workers make an anionic distinction, referring to lakes containing predominantly sulfate as bitter lakes and those containing predominantly carbonate as alkali or soda lakes (Bateman, 1950). Sodium sulfate in its natural form is found in two principal minerals, mirabilite and thenardite. Mirabilite, the hydrous form, is commonly called Glauber's salt. It was discovered by the German chemist, J.R. Glauber (1603-1668), who derived its name from the Latin, sal mirabile, meaning wonderful salt. Thenardite, the anhydrous variety, was named for the French chemist, Louis Jacques Thenard (1777-1857) of the University of Paris (Mitchell, 1979). The largest quantities occur in the form of mirabilite. Sodium sulfate is found in varying degrees of purity, from theoretically pure efflorescent crystals of mirabilite to combinations and admixtures of other salts and impurities. It is a common constituent of some brines; from this source much is extracted commercially. Sodium sulfate also is found in compounds, such as the minerals glauberite, the double salt of anhydrous sodium and calcium sulfate, bloedite, the hydrous double-salt of sodium sulfate and magnesium sulfate, and burkeite, the anhydrous double-salt of sodium carbonate and sodium sulfate. Over 40 minerals contain sodium sulfate in varying proportions; many are of special interest because of their frequent occurrence. Table 1 lists some sodium-sulfate bearing minerals. The reader is referred to other publications (Cole, 1926, Dana, 1932, Grabau, 1920, and Dietrich, 1969) for descriptions of these minerals. Only mirabilite and thenardite will be described herein. Mirabilite Na$04.10H20, contains 55.9% water of crystallization. It is noted for its efflorescence or spontaneous loss of water. On dehydration it changes to the anhydrous form, Na2S0,. Mirabilite is an opaque to colorless, water-soluble mineral that tastes first cool, then slightly bitter. It has a specific gravity of 1.48. It frequently forms as efflorescent, needlelike monoclinic crystals, but generally is found in the massive form. Thenardite, the anhydrous mineral, Na2S04, contains 43.68% Na20 and 56.32% SO3. It ranges from colorless to white and may be tinted shades of gray or brown. It is a water soluble mineral with a slightly salty taste. Its specific gravity (2.67) and hardness (2.5 to 3) exceed those of mirabilite. It commonly occurs in the massive form without visible crystals. Its crystals are frequently tabular pyramids of the orthorhombic system. Sodium sulfate also occurs as a heptahydrate, containing seven molecules of water, but this is unstable and has not been found in the natural environment. The solubility of sodium sulfate has an important effect on the crystallization of the salt in nature, as well as in its production. Its solubility in water generally increases as a nonlinear function of temperature. Below 1.2°C, ice and mirabilite form. As the temperature is increased above O°C, increasing amounts of sodium sulfate become soluble. At 32.4°C, a transition point on the solubility curve is reached, as the decahydrate melts in its own water of crystallization and the anhydrous form crystallizes. With in- creasing temperatures, solubility decreases somewhat. The presence of other dissolved salts changes the transition temperature and solubility characteristics of sodium sulfate.

Jan 1, 1994

The introduction of nitro-carbo-nitrate (NCN) blasting agents has been one of the most significant innovations in the explosive industry in the last twenty years. The most commonly used blasting agent is a rather narrowly proportioned mixture by weight of 94% ammonium nitrate (AN) and 6% No. 2 diesel fuel oil (FO) that is generally designated as ANFO. Other formulations have been developed but none of these have been as extensively used as ANFO. The relative safety of the nitro-carbo-nitrate (NCN) blasting agents almost immediately suggested bulk loading techniques to obtain economies in materials handling. The first attempt at bulk handling was to pneumatically load boreholes with ANFO. When the initial experiments were made in 1957, it was apparent that there were three problem areas which needed study to minimize potential hazards. These were: 1. Static Electricity 2. Fumes 3. Safe Handling Procedures Since those early days, considerable work has been performed in evaluating these problem areas and in arriving at procedures and recommendations that would provide the required safety. Despite the relative insensitivity of ANFO type products, all normal practices for the safe handling of explosives should be obselved. Smoking should not be tolerated in the vicinity of blasting agents and misfires should be handled in the same manner as a misfire with a dynamite charge. The purpose of this Technical Report is to discuss the problems of static electricity and fumes and the solutions that have been developed. Safe handling procedures have been discussed by the Bureau of Mines and are specified by various state and local regulations. Since regulations vary from state to state, the local, applicable regulations must be followed within the area of their jurisdiction. General information is contained in a U. S. Bureau of Mines report, IC 8179, "Safety Recommendations for Sensitized Ammonium Nitrate Blasting Agents" (1963).

Jan 1, 1973

By Rudolf Fahn

Bentonites are best suited for the preparation of acid activated clays with efficient adsorption properties. They are weathering products of acid volcanic ashes and their natural deposits are world-wide spread. The main constituent of bentonite is the aluminium hydrosilicate nontmorillonite with a high ability to swell. Süd-Chemie owns bentonite deposits in Germany and subsidiaries in Italy, Mexico and Arizona.

Jan 1, 1979

By E. M. Schern

The Morenci/Metcalf porphyry copper mining operation, located in southeastern Arizona, currently processes approximately 100,000 tons of ore per day and has copper ore reserves estimated to be in excess of one billion tons. Functional groups within the Geology Depart¬ment (currently staffed with six professional and two support personnel) include mine geology, project geology, and hydrogeology. The primary task of mine geologists is daily preblast bench face mapping with grade and mill amenability estimates. Auxiliary duties include detailed in-pit metal distribution evaluation, mineragraphic work and flux quarry geology. Project geologists conduct explora¬tion and preliminary development projects both within and outside the Morenci district and provide support to ore reserve block modeling and geotechnical work performed by the Morenci mine engineering group. Hydrogeologists handle exploration for and development of groundwater' resources and groundwater flow modeling, mine dump leach solution penetration and recovery evaluation, and tailing dam water reclamation.

Jan 1, 1983

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By Sarkis G. Ampian

This late 1979 Bureau of Mines report presents current stocks and monthly consumption rates for Guyanese refractory bauxite (GRB), with emphasis on the impact of the GRB shortage on the refractory and downstream industries, general background material, and an evaluation of the situation resulting from the strike of Guyanese bauxite workers during the summer of 1979. A study by the Bureau indicated that industry stocks contained little more than a month's supply of GRB--30,000 calcined short tons (CST), at present consumption rates of 25,000 CST per month. The report was requested by the Federal Emergency Management Agency (FW) in response to refractory industry requests that the Government release GRB from the National Stockpile to minimize forecasted industry problems. The report traces the historical innovations in the high-alumina and/or bauxite-type refractory products industry, eventually leading up to fabrication of modern, high-quality firebrick and monolithic refractories.

Jan 1, 1980

By S. Budirsky

J.B. Gwiazda's article deals with an interesting problem that has not been studied thoroughly up to now. Gwiazda has proposed a technical solution that eliminates horizontal load imposed by the supports on the roof. That load is due to a discrepancy between the trajectories of the canopy and the overlying strata. It is true that the improved lemniscate guidance system proposed by Gwiazda could eliminate the horizontal load but, on the other hand, it would represent a higher cost of the supports, and it would complicate the already complicated mechanism of powered supports. That is why we must be convinced that the proposed improvement is actually needed. From that point of view, the following questions should be elucidated. As shown in Fig. 1, the point O travels along the curve a, which produces horizontal load on the roof in the direction of the coal face. As stated by Gwiazda, "such a high load is capable of destroying the roof above the support, causing rock debris to be scattered around the face." The results of our measurements carried out in coal mines of the Czechoslovak part of the Upper Silesian Coal Basin show that the convergence in the modern type of powered supports does not exceed several millimeters (not more than 10) per hour. Similar conclusions were drawn from field measurements performed in the Polish part of the Upper Silesian Coal Basin (see S. Romanowicz and H. Szopka in Proceedings, Scientific and Technical Symposium Simmex '85, Katowice, Poland, pp. 73 -83). The small vertical closure of the supports also induces a small horizontal displacement of the canopy. In addition to this, the transmission of forces between the roof and the canopy is influenced by the layer of debris on it, i.e., the canopy slides without imposing a greater horizontal load on the roof. This was deduced from field measurements of Voest-Alpine F 4/4600 powered supports equipped with rams instead of the common front lemniscate links (denoted as 1 in Fig. 5). In these supports, the pressurizing of the space of piston rod (la in Fig. 5) causes the rams to close, which induces horizontal load on the roof in a direction toward the coal face. After the setting of a support unit, the pressure in the piston rod space was found to drop, frequently to zero, because the small horizontal sliding between the canopy and the debris or inside the debris caused the horizontal reaction from the roof to fall to zero, although the debris were pressurized by vertical setting load. As a consequence, there was no horizontal load imposed by the canopy on the roof. A rise of the horizontal load was recorded only after the passage of the shearer and during lowering of the adjacent units. Thorough information on the results of the afore- mentioned field measurement is given in the paper, "Analysis of the performance of shield powered supports installed in a thick seam," published in Mechanizacja i automatyzacja gornictwa, 1982, No. 12, pp. 38-46. An English translation of the paper is available from the author. We have deduced from our observations that horizontal load imposed by the canopy on the roof in the direction toward the coal face is useful for roof control because it limits the displacement of the immediate roof toward the gob particularly during the rise working. We came to the conclusion that horizontal load on the roof toward the coal face should be induced on purpose. Could Mr. Gwiazda prove with the results of field measurements his contrary opinion? Has the small horizontal displacement of the canopy toward the face actually had an adverse effect on the roof? As shown in Fig. 2, during the setting of a support unit, the canopy imposes a horizontal load on the roof in the direction of the gob. In this case, I share Gwiazda's opinion that the horizontal load could have an adverse effect on the roof. Nevertheless, two important circumstances were neglected - the horizontal compliance of the supports and the interaction of adjacent units. The question of the compliance was analyzed precisely by I. Krumnacker in Gluckauf -Forschungshefle, 1984, NO. 5, pp. 219-223. He found that due to the clearance between the hinge pins and the eyes of the sheild and due to the elasticity of the steel structure,

Jan 1, 1988

By Elza Burch

Introduction The feeding to zero concept involves feeding 600 µm x 0 (28 mesh x 0) size raw coal to heavy media (magnetite) cyclones along with the +600 µm (+28 mesh) size coal. Traditional circuits employ desliming or removing the 600 µm x 0 (28 mesh x 0) size fraction and feeding the cyclones +600 µm (+28 mesh) size coal. The feeding to zero concept recirculates 600 µm x 0 (28 mesh x 0) fines in the circuit. At the same time, a portion of the fine material is continuously withdrawn and recovered. This, in turn, prevents a fines buildup. This concept eliminates desliming screens and secondary fines circuitry for recovery of 600 x 150 µm (28 mesh x 100 mesh) coal. The result is a very simple circuit. Feeding to zero at Island Creek Island Creek Corp. was the first involved with the new concept in 1976. The company needed a temporary plant for the 9.5 mm x 0 (0.4 in. x 0) raw coal at its Pond Fork mine, near Madison, WV, while a full-scale plant was being designed and built. At that time, the Childress Corp., of Beckley, WV, became interested in the feeding to zero concept. Island Creek awarded a contract to Childress to build a single cyclone modular plant, incorporating this feeding to zero concept. The plant was erected in three months. The Pond Fork modular plant proved successful in attaining the desired feed rate of about 63.5 t/h (70 stph), while maintaining good separating efficiencies and low magnetite consumption rates. The 9.5 mm x 150 µm (0.4 in. x 100 mesh) clean coal was recovered and the 150 µm x 0 (100 mesh x 0) size was disposed of to waste. The full-scale plant was completed about two years later and the Pond Fork modular plant was moved to Holden, WV. There, it was incorporated into the Holden 29 preparation plant as a separate circuit for cleaning -25 mm (-1 in.) coal. In 1976, a similar plant was installed in Virginia by another company. These two plants are believed to be the first two operational plants in the United States incorporating the feeding to zero concept. Island Creek subsequently contracted with Childress for an identical plant at the Coal Mountain operation in West Virginia. The plant operated for three years before the mine was closed. The unit was then moved to the Spurlock mine near Martin, KY where it continues to operate. The successful operation of the Pond Fork and Coal Mountain plants before and after relocation proved both the performance and moveability of this type of circuit when constructed in a modular fashion. Since the first two plants were built, Island Creek has incorporated the feeding to zero circuit in nine additional plants. A grand total of 33 cyclones have been installed using this concept. One is the Providence mine, near Providence. Providence preparation plant Island Creek contracted with J.O. Lively Corp. of Glen White, WV in July 1978 for the construction of the Providence preparation plant. The plant began operation in February 1979. The construction period was about halved by building the plant with modular design concepts. Prefabricated sections, floors, and sides were brought in as units and then bolted in place. The Providence plant has a good track record of processing coal at a feed rate of 454 kt/h (500 stph). Feed coal to the plant has an average ash content of about 18% and sulfur content of about 4.5%. It contains about 22% refuse. The coal product has an average ash content of about 8% and sulfur content is about 3%. Raw West Kentucky No. 9 seam coal is conveyed from a box cut in the Providence mine to a rotary breaker. The breaker is fitted with 74 mm-diam (3 in.-diam) opening breaker plates. Therefore, it is well suited for removing trash, roof bolts, wood, and pyritic balls that are common in Illinois Basin coal. The -75 mm (-3 in.) coal is conveyed from the rotary breaker to

Jan 8, 1987

Brace yourself-what used to be a 1 billion ton deposit is now a 1 Gt deposit. And that 1520-hp ball mill is now a 1.134-MW ball mill. Beginning with this issue, all engineering units and statistics in MINING ENGINEERING will be reported in SI units. The exception to this is the Technical Papers section, where units may still be reported by dual notation. For the past four years, ME has reported all units of measurement first in SI, followed in parentheses by the equivalent English value. This metrication policy, adopted by the SME-AIME Board of Directors in September 1975, was designed to aid magazine readers through a transitional period from English unit reporting to SI unit reporting. In October 1979, the SME-AIME Publications Board completed the English-to-SI transition by requesting the editorial staff, beginning February 1980, to report exclusively in SI. The Publications Board also requested the editors to prepare a chart of SI conversions useful to ME readers; this information follows on pages 156-157. The What and Wherefore of SI In 1960 the General Conference on Weights and Measures met in Sevres, France, and recommended that a new metric system-one that could uniformly cross all international borders-be adopted to replace all other metric systems. The outcome of this recommendation was the SI system of units, an abbreviation for le Systeme International d'Unites (the International System of Units). It replaces the more familiar cgs (centimeter-gram-second) system and the m/ksA (meter-kilogram-second-Ampere) system. SI consists of seven base units, two supplementary units, and a number of derived units. Other units, called allowable units, may be used even though they are not part of SI. The seven base units are: meter (length); second (time); kilogram (mass); mole (amount of substance); kelvin (temperature); ampere (electric current); and candela (luminous intensity). The two supplementary units are radian, a measurement of plane angle, and steradian, a measurement of solid angle. Supplementary units can be used in combination with SI units, e.g., radiant intensity is expressed in W/sr, watts per steradian. Derived units are combinations of base and supplementary units. Velocity is expressed as meters per second (m/s), and acceleration as m/s'. Using the relation established by Newton's second law of motion, force is expressed as kg-m/s2, more commonly known as the newton (N). Energy (work) is force times distance, expressed as kg-m/s2 and known as the joule (J). Power is work per unit time, expressed as kg, m2/s3 and known as the watt (W). Allowable units are the exception to the rule-they are not taken from base or supplementary units but, nevertheless, are used in the SI system. Volume units are normally expressed as the cubic measure of a length, such as cm 3. But a fluid volume of a cubic decimeter (dm3) is called a liter (L); the "L" is capitalized in another special SI rule to keep the lowercase "1" from being confused with a numeric "1". Other allowable units are degrees Celsius (°C) for temperature; and the minute (min), hour (h), day (d), and year (a) for time measurements. The prefixes in the table on the following page are used with SI units to express the magnitude of a measured quantity. A mill that processes 1000 tons of ore per day would be designated a 1 kt/d mill; a mine producing 10 million tons of ore per year would be called a 10 Mt/a mine. The rule here is to use prefixes that keep the value in a convenient range from 0.1 to 1000. In other words, it would be incorrect to write 25 000 kg-the prefix should be changed to indicate 25 Mt. One of the distinct advantages of SI is that it retains the decimal relation between multiples and submultiples of the base units for each physical quantity except for time. One more rule of SI syntax-numbers are separated into groups of three with a space between each group rather than a comma, e.g., 3,839,174 becomes 3 839 174. By the same token, a decimal number is also broken by spaces, e.g. 0.738153 is written as 0.738 153.

Jan 2, 1980

By Dennis S. Kostick

Introduction Soda ash is known chemically as sodium carbonate, an important inorganic chemical. It has been produced for several centuries by processing certain vegetation and minerals. The US soda ash industry has evolved from several small sodium carbonate mining operations in the West. Now, a nucleus of six companies produce about one-fourth of the world's annual soda ash output US producers currently dominate the world market. But certain international events are occurring that will reshape the domestic soda ash industry in the next decade. Historical perspective Soda ash is used mainly in the manufacture of glass, soap, dyes and pigments, textiles, and other chemical preparations. All of these are the first basic consumer products produced by developing societies. About 3500 BC, the Egyptians became the first society to use crude soda ash. The soda ash was used to make glass containers. It was most likely obtained from dried mineral incrustations around alkaline lakes. Soda deposits were virtually nonexistent in western Europe. So people resorted to burning seaweed to obtain the ashes. The ashes were then leached with hot water and the solute was recovered after evaporating the solution to dryness. The solute, a crude "soda ash" was impure. But, it could be used to make glass and soap. These two products and industries were important to the population and economic growth of the region. About 11.5 t (13 st) of seaweed ash was required to produce about 0.9 t (1 st) of soda ash. Along the coasts of England, France, and Spain, seaweeds with varying alkali contents became important items of commerce and sources of soda ash before the 18th century. The LeBlanc process used salt, sulfuric acid, coal, and limestone. It became the major method of production from about 1823 to 1885. In the early 1860s, Ernest and Alfred Solvay, two Belgian brothers, successfully commercialized an ammonia-soda process to synthesize soda ash. It used salt, coke, limestone, and ammonia. The Solvay process produced a better quality product than the LeBlanc method. In 1879, Oswald J. Heinrich presented to the Baltimore meeting of AIME, a paper entitled "The manufacture of soda by the ammonia process." The paper compared the two processes and foretold the demise of the LeBlanc technique. World production of soda ash in 1880 was 680 kt (750,000 st). Of that, 544 kt (600,000 st) was produced by the LeBlanc process. Of the 2.8 Mt (3.1 million st) of soda ash produced worldwide in 1913, only about 50 kt (55,000 st) was by the LeBlanc method. The LeBlanc process was never used successfully in the US, except for a brief period from July 1884 to January 1885 in Laramie, WY. Previously, soda ash had been produced by burning certain plants, as exemplified by the early Jamestown colonists, or by recovering small quantities of natural sodium carbonate found in alkaline lakes, such as those found near Fallon, NV, and Independence Rock, WY. Before the 1884 startup of the first synthetic soda ash plant in the US at Syracuse, NY, most of the domestic soda ash demand in the East was met by imports, primarily from England. Large-scale commercial production of natural soda ash began in California in 1887 from surface crystalline material at Owens Lake. Production from sodium carbonate-bearing brines at Searles Lake began in 1927 (Fig. 1). In 1938, during exploration for oil and gas in southwestern Wyoming, a massive buried trona deposit, presumably the world's largest, was accidentally discovered. Recent mineral resource evaluation by the US Geological Survey and the US Bureau of Mines indicates that the Wyoming trona deposit contains 86 Gt (93 billion st) of identified trona resource in beds 1.2 m (4 ft) thick or greater. Additionally, there is about 61 Gt (67 billion st) of reserve base trona. Of this 36 Gt (40 billion st) is in halite-free trona beds and 24 Gt (27 billion st) is in mixed trona and halite beds. In 1953, the Food Machinery and Chemical Corp. (later shortened to FMC Corp.) became the first company to mine trona in Wyoming. Soda ash demand increased.

Jan 10, 1985

By Charles S. Robinson

The Henderson ore body is east of the Continental Divide in the Front Range of the Colorado Rocky Mountains about 50 miles west of Denver. The ore body is being developed for mining by workings that extend more than 3500 feet below the topographic surface. During construction of the mine workings, ground water in fractures in crystalline rocks has been encountered. From 1969 to 1973, Charles S. Robinson & Associates, Inc., in cooperation with the engineering and geologic staffs of the Climax Molybdenum Company, a division of Amax Inc., studied the ground water geology of the Henderson mine development area. Detailed geologic records of the Henderson mine were available as a result of the exploration, evaluation and development of the mine. Records of the flow, chemical composition and temperature of the ground water at and near the surface, from the entire mine, from portions of the mine, and from exploration holes drilled in the mine, were kept as the mine was developed. Where possible, ground water pressures in drill holes and changes in pressure with time, were recorded. The geologic, hydrologic and chemical data were correlated. The following conclusions were drawn from the correlations: The ground water in the Henderson mine occurs in fractures in Precambrian igneous and metasedimentary rocks and in a series of Tertiary intrusive rocks. Two zones of ground water were recognized; an active zone, in which the fracture systems are recharged and discharged annually into the surface water system, and a passive zone, in which the fractures were filled with water in the geologic past. Discharge from the passive zone is through mine workings that intersect the fracture systems. Discharge from a fracture or fracture system is initially high but decreased with tire!. The rate of decrease in flow (or pressure) is exponential. By measuring the change in flow with time, it is possible to calculate the amount of flow at some given period of time in the future. Based on changes in rate of flow and pressure, and the geology, and assuming rates of mining of the ore, it will be possible to predict the volumes of water that will have to be handled during the life of the mine and the area around the mine that will be drained by the mining operation.

Jan 1, 1976

By Charles B. Escapule

The first successful heap leach of silver ores in the State of Arizona has been under continuous operation since early 1976 at the State of Maine Mine, two miles southwest of Tombstone, Arizona. The silver produced from leaching old mine dumps at this property vas the first in the Tombstone mining district to achieve a 999+ purity. While cyanide leaching of silver and gold ores is a time-proven accepted practice, a relatively new method of treatment called "heap" leaching was employed. Heap leaching is a comparatively inexpensive fast way to recover precious metals from low grade material.

Jan 1, 1981

By Henry N. McCarl, Gary Waters, William A. Vogely

INTRODUCTION In the past three decades, world trade has grown from less than $100 billion per year in the years following World War 11 to slightly more than $300 billion in 1970 and over $2,000 billion in the early 1980s. Adjusted for inflation, the real growth of world trade in 1972 dollars increased from less than $175 billion per year in the early 1950s to slightly more than $342 billion in 1970 and almost $700 billion per year by the early 1980s. Thus trade as a whole has increased a little more than twofold in real terms in the past decade." Mineral commodities constituted approximately 20% of the value of total world trade in 1970, and significantly more than that proportion in terms of physical volume. By 1980, de- spite rapidly rising petroleum prices, mineral commodities accounted for roughly 18% of the total value of world trade. Crude oil alone ac- counted for about 16%, making it the single most important commodity in international trade. The worldwide recession of 198 1- 1983 brought a dramatic decline in both volume and price of crude oil, shrinking the impact of mineral commodities to roughly 10% of the value of total exports, worldwide, in 1982- 1983. The decline of mineral commodities as a pro- portion of the value of world trade in the decade 1970- 1980 is due largely to the increasing importance of manufactured goods and consumer products such as automobiles and electronic items. During this period, mineral commodity trade expanded roughly sixfold in terms of current dollars to an estimated $350 to 360 billion in 1980. This was equivalent to doubling the real value of mineral commodities in international trade in the 1970- 1980 period in constant dollars. The U.S. Bureau of Mines has expanded the traditional definition of mineral commodities to include semifinished products such as iron and steel (in addition to iron ore), refined metals such as copper, lead, zinc, aluminum, and tin, concentrates, scrap, and petroleum products. Since large quantities of essentially mineral derivatives travel in this form, it is reasonable to use these expanded definitions in stating the role of minerals in world trade. Thus minerals in world trade comprise 30 to 33% of total export trade in all commodities during the period 1976- 1980 (see Table 1.3.1). In addition, the U.S. Bureau of Mines estimates the value of all mineral commodities consumed in the world (at the national level) and compares this value to total worldwide commodity consumption. By this system, mineral commodities range from 30 to 40% of all commodities consumed each year in all countries during the period 1976- 1980 (see Table 1.3.2). The growth in export trade in mineral commodities by commodity group is shown for the period 1976-1980 in Table 1.3.3. These data parallel the swings in worldwide economic activity during the same time period. By any of the above measures, minerals, especially the fossil fuels, constitute a significant proportion of both total economic activity and

Jan 1, 1985

By S. Fernández, L. Klinger, T. Vallejo

In the municipality of Rio Quito, department of Chocó, Colombia, alluvial gold has been illegally exploited by suction dredging. The project consists of using geospatial techniques to identify degraded areas and analyze the viability of remediation dredging to mitigate environmental liabilities. The area has sediments contaminated by mercury and cyanide, so it is necessary to reconform the riverbed with the remediated material. The work is part of the Mine Closure Plan that seeks to jointly articulate the communities, academia, the company, and the state

Mar 2, 2022

By John L. Halls

INTRODUCTION In this section, it is assumed that the reader is an engineer with sufficient training in underground mining work, preferably with budgetary control, to understand the mining methods outlined in other sections of this handbook. The mining-method descriptions are not repeated herein; only sufficient information is provided to enable the reader to comprehend the development of the costs presented. All costs are in 1977 US dollars, except where explicitly stated otherwise. CAPITAL INVESTMENT The capital investment in the underground mining portion of any mining property includes the following: 1) The development of original access involving adits, shafts, declines, etc. It includes the necessary equipment for transporting men, materials, and rock. 2) The purchase of mine equipment including loaders, underground haulage equipment, fans, pumps, electrical distribution, surface facilities associated with the underground mine, etc. 3) Sufficient preproduction mine development to assure steadily increasing production after startup. De- pending upon the mining method used, the preproduction development should provide a supply of three months to one year of ore that is developed and is ready to be stoped. Underground development also includes excavations other than shaft stations, loading pockets, sumps, etc., which usually are associated with the means of access and are excavated by the contractor. These excavations could include a crusher cavern, an ore- storage pocket, and underground workshops that may or may not be included by the shaft-sinking contractor. Portions of these could be delayed until after production begins, but it is preferable to complete them during the development phase, so there is no interference with the production buildup. 4) The staff and labor costs for all phases from inception through final production. 5) The provision of mine-surface equipment and facilities strictly applicable to the underground portion of the operation. The required facilities include a changehouse with a lamp room, a first-aid station, compressors, an electrical substation, workshops, a ware- house, water pipelines to the concentrator, water storage tanks, explosive storage, etc. A concrete-batch plant, a sand-fill plant, and boilers for heating may be required. 6) Other direct and indirect costs. 7) Engineering, design, and procurement costs. 8) Contingency allocations. 9) Contractor's fees, startup costs, and owner's costs. Original Access As part of a study for the US Bureau of Mines, Dravo Corp. conducted an extensive literature research survey covering underground operations in the United States between 1963 and 1973. Table 1 is compiled from the section of that study pertaining to the types of access to underground mines. Table 1 does not include distances attributable to the deepening of existing shafts, nor does it include shafts bored by raise drills or downhole machines that need existing underground workings for the disposal of the cuttings. During the period under review in the report ( 1963 through 1973). concrete-lined circular vertical shafts were the predominant means of providing the first access to a new underground mine; these shafts accounted for 80% of the original access in terms of the number of accesses completed, and 75% in terms of the total verti- cal linear distances. The inside diameters of the conventional circular shafts ranged from 3.7 to 7.0 m (12 to 23 ft). Most (72% ) of these conventional shafts were for production,

Jan 1, 1982

By Kerry J. Stanaway, Thomas E. Garnar

Titanium was discovered in 1791 and, four years later, was named after the Titans of Greek mythology. Titanium metal was first made in 1910, but was not used commercially until the 1950s when it was nick-named The Wonder Metal of the Age. The first titanium dioxide (TiO2) white pigments were produced commercially in 1915; more than 90% of the titanium minerals produced in the world today are used in the manufacture of white pigments. The remaining 10% are used to manufacture titanium metal and other uses such as welding rod fluxes. Titanium mineral mining began in the United States between 1880 and 1900 in Chester County, Pennyslvania. Rutile from a deposit near Roseland, VA, was mined in 1901. The use for the rutile was limited to production of a few titanium chemicals and for coloring ceramics. Larger scale mining for ilmenite, rutile, and zircon began in Florida to provide chemical materials for use in World War I in 1918. This mining ended in 1928. Large scale mining of titanium minerals in Australia began when Benick, Anderson, and Porter (later called Tamco) started mining the beach sands at Yamba in New South Wales in 1935. Ilmenite production from beach sands in the State of Kerala in southwestern India also began in the 1930s. The 1940s saw many changes in titanium mineral mining. Extensive deposits of ilmenite and ilmenite-bearing iron ores were discovered in eastern Quebec. The Pigments Department of E.I. Du Pont de Nemours & Co. lost their high grade Indian ilmenite supply after World War II began, and they were forced to use lower grade ilmenite from National Lead's (now NL Industries) Sanford Lake Deposit, near Tahawus, NY, in the Adirondack Mountains. At the end of the war, Du Pont undertook an intense search for a domestic ilmenite source. With the cooperation of the US Bureau of Mines, Du Pont geologists found a very large deposit east of Starke, FL, in 1946. In 1948 Du Pont entered into a contract with Humphreys Gold Corp., who operated a heavy minerals mine near Jacksonville, FL, to construct a plant and operate it for a period of ten years. The plant started up in 1949 and operated under Humphreys' management until the contract expired in 1958. Du Pont has operated the plant since that time. In 1956, a second plant was designed and built by Du Pont a few kilometers northeast of the town of Lawtey. The 1960s saw titanium mineral production expand in the southeastern US with Humphreys' contract mining Du Pont's ore body at Folkston, GA. In the early 1970s. American Cyanamid and Union Camp Paper Co. formed a joint venture company, Titanium Enterprises (TE), to mine and separate titanium minerals from Union Camp's ore deposit southwest of Green Cove Springs, FL. In 1975 the TE plant was shutdown and later sold to Associated Minerals (USA), Inc. (AMC), a subsidiary of Renison Goldfields Consolidated Ltd., an Australian company. It now operates under the name RGC (USA). Mineral production grew in the 1970s with the development of extensive heavy mineral deposits at Eneabba in Western Australia and Richards Bay in the Republic of South Africa. The 1980s saw the construction of a very large ilmenite mine and synthetic rutile processing plant in the state of Orissa southwest of Calcutta, India. At the same time, concern for titanium mineral shortages expected by the end of this century, gave rise to a flurry of exploration activity during the mid 1980s. As a result, large deposits were found in Virginia and North Carolina. New deposits were also discovered in Australia. Fig. 1 shows the distribution of world titanium deposits and mines. The world TiO, production capacities are given in Table 1.

Jan 1, 1994

By J. M. Applegate, C. L. Morgan

Introduction In the same way that a blueprint communicates informa¬tion about a structure, a systems architecture should provide information that guides the construction and implementation of a computing system. Taking the blueprint analogy a step further, a plot plan can also show where, in the "bigger picture," a particular design fits (if in fact it does) and what a builder needs to know about the environment in which he or she will operate. In the context of business systems, a meaningful architecture should convey the following: • the primary purpose of computer-based systems rela¬tive to particular tasks or operations, • the relationship of the systems to each other and an indication of their interaction and dependencies, • the relationship of the systems relative to the organiza¬tion and its objectives, • indications of where IS/IT resources can best be allo¬cated to support the organization and • the overall context in which new systems and technolo¬gies can be assessed for both their impacts on current systems and on the organization itself. In theory, a well-designed systems architecture should allow an organization to focus its resources on those aspects of its business that matter most, thus, leveraging investments in information technology to a competitive advantage. In addition, the way that a computing systems architecture is represented should be un¬derstandable to non-IT staff, and it should be eas¬ily recognized by the people it supports. This pa¬per attempts to show that a meaningful systems model exists and that its applica¬tion to mining can achieve the desired results. The following discus¬sion is based on the au¬thors interpretation of a concept referred to as the computer integrated manu¬acturing (CIM) model, as developed by Allen-Bradley. An inherent strength of the CIM approach is that the model combines aspects of business-reporting structures with the manufacturing process itself. What results is a systems frame¬work that is aligned with relevant business processes and is, therefore, easier for non-IS/IT personnel to interpret. The successful application of the CIM model can be seen in a large number of both discrete and process-manufacturing facilities throughout the United States. This paper provides a brief overview of the CIM model and then describes how the model can be applied at a conceptual level to the mining industry. Finally, this paper proposes an extension to the CIM model that the authors believe will make it easier to understand and apply. CIM model As shown on Fig. 1, the CIM model begins with the most detailed levels of process equipment and works its way up to the highest level of consolidated financial reporting. Systems at each level are characterized by common attributes, users, time frames, etc., providing a clear picture of their interac¬tion. Further analysis of the relationships between systems and between levels provides insight into the opportunities for integration. Regardless of the industry or processes involved, the CIM model provides a framework in which the characteristics of systems and the data they manage can be predicted based on the nature of the business function they support. In general, as data moves upward from the lower rungs of the CIM model, it changes from de¬scribing the physical na¬ture of a process to describ¬ing its financial aspects (e.g., from productivity to cost). The time frame to which data applies also transitions from a very short duration, perhaps only milliseconds, to financial or regulatory reporting time frames that may span years. In a simi¬

Jan 1, 2000

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The reliable quality assessment of grouting operations is one of the key issues in establishing impermeable grout curtains around shafts, drifts and tunnels. Reliable opera¬tional information on the decrease in the permeability of rock during the grouting process facilitates the assessment of the effectiveness of the completed curtain and assures production quality. Quality control and assurance also fa¬cilitates the early detection and elimination of possible de¬fects in the grout curtain. Thorough control of the grouting process provides the capability to evaluate decreases in the permeability of water-producing fracture zones prior to the beginning or the continuation of excavation of the shaft, tunnel or drift. In order to achieve this objective, post¬grouting hydrodynamic investigations are conducted prior to the initiation or re-initiation of work on the underground workings to measure the decrease in the permeability coef¬ficient of aquifers during grouting operations. 8.1 QUALITY ASSURANCE AND CONTROL OF THE ISOLATION OF ROCK BY GROUTING THROUGH HOLES DRILLED FROM THE SURFACE Quality assurance and control are based on measuring the decrease in the permeability coefficient of water-pro¬ducing horizons as a result of injecting grout into the drill¬holes. For this purpose, the value of the acceptable residual permeability coefficient is calculated beforehand based on the acceptable value of ground water inflow into the shaft, drift or tunnel as specified by the "construction norm and integrated practice (SNIP)" (Anon., 1976). Having ob¬tained the values of aquifer thickness, permeability, and head distribution for each aquifer from flowmetric data (dis¬cussed in Chapter 3), the average value of the permeability coefficient of all aquifers intersected by the shaft is calcu¬lated as [ ] where Ki is the permeability coefficient of the "ith" aqui¬fer; Pi is the pressure of the "ith" aquifer; and M; is the thickness of the "ith" aquifer. The total expected inflow of ground water into the shaft is [ ] where RK is the radius of the contour of influence; RCFB is the radius of the shaft; and p is the coefficient of dynamic viscosity of the ground water. The necessary reduction of the coefficient of permeabil¬ity of the aquifer [ ] where Q is the expected inflow of ground water into the shaft from eq. 8.2; Q.-fin is the acceptable inflow of ground water; and a is the safety factor. By knowing the average permeability value of the rock and the permeability reduction coefficient, it is possible to determine the value of the acceptable residual permeability coefficient for all the aquifers [ ] The values of the permeability coefficient, pressure, and thickness for each aquifer are determined from the data gathered during the hydrodynamic investigations conducted in grout holes prior to beginning the grouting operations as discussed in Chapter 3. After injecting the grout into the first of the holes, the injection into each of the following grout holes is preceded by repeated hydrodynamic investi¬gations. Using the data obtained for each water-producing zone, a graph is constructed where the numbers representing the grout holes in their order of grouting are placed on the abscissa, and the values of the aquifer permeability coeffi¬cient are placed on the ordinate (Fig. 103). This curve char¬acterizes the change in permeability of the given hydro¬stratigraphic unit as grouting operations are completed, one hole at a time. The results of grouting operations are considered to be positive when each sequential observation records a steady decrease in the permeability of the aquifer. The permeabil¬ity measured prior to the injection of the grout into the last hole must be lower than or equal to the acceptable perme¬ability as determined by the application of eqs. 8.1 through 8.4. Table 8.1 presents illustrative data on the hydrogeo¬logic properties of aquifers at a mine shaft prior to the initiation of grouting operations. Data also are presented on the reduction in the permeability coefficient during the in¬jection of the grout. The shaft in question intersects eleven aquifers. Based on data concerning aquifer pressure, thick¬ness, permeability, and the expected inflow rate of water for each of the strata, the average permeability coefficient is calculated according to eq. 8.1 as K = 0.2 x 10-12 m2. The

Jan 1, 1993

By William R. Wayment, Wayne S. Cusitar

INTRODUCTION The process of mining removes valuable and eco¬nomically recoverable minerals from the earth. In the case of hard-rock mining, this process involves breaking and removing ore, while leaving the waste rock or host rock as intact as possible. Prior to mining, ground stresses exist in the ore and the host rock as a result of weight and tectonic forces; for the purposes of this chapter, the ore is considered to support the host rock on both the hanging and footwall sides. Two basic phenomena accompany mining. First, the hanging walls and footwalls tend to move together as the ore is removed. Second, high stress gradients de¬velop in the rock periphery of the opening; the magni¬tudes of the stress gradients vary widely, depending upon the local geometry and conditions. At various times, mining methods have used square¬set timbers, waste-rock backfill, alluvial sand fill, and tailings backfill materials, with or without cement, to replace the loss of support caused by the removal of the ore. Cut-and-fill mining has evolved to use hydraulically placed tailings for the backfill. In the stope, this fill provides massive support to the walls, thereby reducing stresses in the back. It also provides a floor from which men and machines can work to drill and blast, to haul ore, and to scale and bolt the back. In many cases, the addition of cement provides a competent free-standing wall when the fill is exposed during mining of adjacent stopes or during pillar recovery operations. In most cases, the use of fill does not eliminate standard tem¬porary or local roof control measures such as roof bolts and wire mesh. Fill Characteristics It is important that backfill used for ground support be of maximum stiffness in relation to wall closure. The stiffness of the fill is related to its placement bulk density or void ratio. In the case of cemented tailings, this is enhanced by the cement bonds. During stope closure, individual fill particles experience rotation and crushing at point contacts, but they are subject to little relative translation. The stiffness approaches infinity as the void ratio approaches zero, until the ground forces reach equilibrium, and closure stops. The gradation of particle sizes and minimal segregation resulting from good hy¬draulic placement both contribute to a high degree of early stiffness in the closure history of classified tailings backfill. This also implies that the fill should be resistant to creeping under high static loads. All mines experience shock loading situations such as blasting, and many mines experience rock bursts that can apply sudden and very high loads to the fill. Certain conditions of void ratio, particle size, and moisture con¬tent can result in mass fluidization of unconsolidated fill during such shock loading. Two key factors in avoid¬ing fluid behavior are the provision of a well-drained fill through desliming or classification and the achievement of a small degree of consolidation through the addition of cement in ratios as lean as 40 to 1 (2.4% cement). The acceptance of classified mill tailings as backfill material has been enhanced by its relative economy, including its source, preparation, delivery, and stope distribution. Transportation from the backfill prepara¬tion plant on the surface into the stope usually can be accomplished with low power consumption, with low labor costs, with few capital costs, and at high tonnage rates. An additional benefit is a reduction of the storage volume required for surface disposal of the tailings stream, with a resultant improvement in the environ¬ment. The costs of placing cemented tailings backfill generally range from $1.65 to $6.61/t ($1.50 to $6.00 per st) of fill depending upon the cement ratio used. Scope This chapter examines a major backfill program at a mine and mill complex where the backfill facilities are included as a part of the original planning. The backfill preparation plant is intended to make maximum use of remote controls and automation. The logistics asso¬ciated with material storage, handling, and metering in the plant are outlined. Important considerations in specifying equipment for typical service conditions are discussed. The fill delivery system is described, includ¬ing the boreholes, level lines, and stope distribution lines. Stope preparations for fill confinement and drain¬age are described, as are the techniques of placement and distribution. Methods of controlling and directing the flow of drainage water and slimes are presented, along with the facilities and equipment for clarification, pumping, and sludge removal. As shown in Fig. 1, the fill program encompasses most aspects of the mining and refining facilities, so the discussion is brief in each area. The preparation of backfill of a uniform and con¬trollable quality is presented here as a material handling problem. The balance of the presentation concentrates on the logic of the system, the costs where meaningful (in 1977 US $ unless otherwise specified), and some operating "tricks" that have contributed to system per¬formance and reliability. SURFACE PREPARATION PLANT A backfill preparation plant is required on the surface. This plant generates suitable fill material from the stream of mill tailings. Among the functions of the plant are: 1) Slimes are removed to improve the percolation characteristics of the fill; typically, particles finer than 325 mesh are removed from the tailings. 2) Tailings and sand storages are provided so that the steady stream of tailings produced by the mill can be accepted. Provision is made to accommodate the cyclic high-volume demand from the mine. 3) The plant stores the cement and contains meter¬ing and mixing facilities to provide fill of a controllable and uniform quality. 4) The hydraulic backfill delivery system is a part

Jan 1, 1982

By K. Hanumantha Rao, K. S. E. Forssberg, H. R. Manouchehri

This paper summarizes the practical considerations of the various electrical separation methods used in mineral processing. The effective factors related to mineral characteristics and equipment attributes are described and analyzed with reference to conventional industrial applications of the electrical methods. Some new aspects of electrical methods application and their potential for mineral beneficiation, apart from the mineral systems that have been successfully beneficiated, are also presented. In addition, wet dielectric separation, which now appears to be promising, is briefly cited.

Jan 1, 2000