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By G. A. Hufford

The cutoff grade applied to a disseminated orebody mined by open pit affects both the waste ratio, hence the cost of producing ore, and the ore grade, hence the quantity of metal recovered. As such it has a significant impact not only on the annual profit or loss, but also on the life of a mining property. Cutoff Grade Strategies Some commonly used objectives in calculating cutoff grades are as follows: Minimize Unit Cost of Product (1) This is found by adding all costs associated with each possible cutoff such as mining, processing, and marketing. This is divided by the quantity of marketable product obtained at each cutoff grade, then finding the minimum. Because the market value of the product does not enter the calculation it has no effect on this strategy: the same cutoff is selected in good times and bad.

Jan 1, 1986

By G. J. Grott

Direct Production Costs for Solar Sa 1t approximate the trucking cost for 160 Km. Operations near markets appear unusually attractive at first look. However, many desert operations have been plagued by dust storms yielding salt too high in insolubles for most uses. This paper relates the history of a successful operation in a dust storm area. Saturated brine from the solution mining of LPG storage caverns is evaporated in solar ponds yielding 99.95% sodium chloride (dry basis). Purchased energy is one-eighth that of vacuum pan operations. Production scale tests have proved it works at another desert location using a more complex brine. A new approach for refining soluble salts at ambient temperatures is demonstrated by bench scale and pilot work. Again, savings in purchased energy exceed 80%, even when temperatures are raised to reduce refining time.

Jan 1, 1988

By A. T. Yu

Bulk sampling is define' as the ?process of extracting a small fraction of material from a large bulk ...sufficiently representative ...for the intended purpose.? (ref. 1). Today's pressing need for simple, effective and reliable sampling systems is intensified by the growing demand for quality control as a means to cope with inflation, competition and rapidly increasing consumption of raw materials. Much has been written on bulk sampling (ref. 2 thru 9). It involves a wide range of subjects cutting across the domains of physics, chemistry, geology, mining, materials handling and processing, as well as probability and statistics. Since the problem itself is complex, any in-depth treatment cannot avoid a somewhat lengthy coverage. Of all of the related topics. Arbitrary, haphazard studies, on the other hand, can leave out important details and obscure or confuse the issue.

Jan 1, 1970

By Cyril W. Draffin

Introduction Americans have long been proud of their ability to recognize a need, define it, and then develop workable solutions. The U.S. underground coal conversion (UCC) program is a good example. In 1974, the need to utilize coal resources that could not be mined was recognized, and Congress gave the Bureau of Mines $1.7 million to gain a better understanding of the UCC technology and its potential applications. Since 1977 when the U.S. Dept. of Energy (DOE) was formed, workable technologies have been tested, the market for underground coal conversion has been recognized, and extensive industrial interest has developed. Perspective The U.S. is blessed with extensive coal resources. Coal supplies about 19% of our national energy needs, yet recently promulgated environmental standards and increasing mining costs have limited direct coal use. UCC potentially can use coal that is economically or technically infeasible to mine because the coal is too thick, deep, dirty, high-ash, and wet; has an excessive angle of dip; or has unpredictable and poor overburden characteristics that make mining unsafe. Coal can be gasified underground or perhaps liquefied below the earth’s surface. Underground coal gasification (UCG) is the prime element of the underground coal conversion program. It enables coal to be converted to gas underground. UCC Coal Resource Recent data show a total U.S. coal resource base of about 6.4 trillion tons. Of this, 1.8 trillion tons might be accessed by UCC, including some minable coal. A commercial UCC process could quadruple roughly the 450 billion tons of U.S. coal reserves, which presently are recoverable by mining. The U.S. coal resources are divided into five sectors. The two most eastern sectors contain relatively thin, swelling bituminous coals that are difficult to gasify. Of the two, the bituminous coal in the Eastern Province is the most difficult because it is deeper than , in the Interior Province and the terrain is more rugged. The sector along the Rocky Mountains contains huge amounts of thick low-rank coals that are the most promising for early development by UCC. The coal has good continuity and the terrain is suitable. Low-rank coals such as subbituminous and lignite are highly reactive and tend to shrink and disaggregate when heated. Most of the coal in the Pacific

Jan 1, 1980

By Cyril W. Draffin

Introduction Americans have long been proud of their ability to recognize a need, define it, and then develop workable solutions. The U.S. underground coal conversion (UCC) program is a good example. In 1974, the need to utilize coal resources that could not be mined was recognized, and Congress gave the Bureau of Mines $1.7 million to gain a better understanding of the UCC technology and its potential applications. Since 1977 when the U.S. Dept. of Energy (DOE) was formed, workable technologies have been tested, the market for underground coal conversion has been recognized, and extensive industrial interest has developed. Perspective The U.S. is blessed with extensive coal resources. Coal supplies about 19% of our national energy needs, yet recently promulgated environmental standards and increasing mining costs have limited direct coal use. UCC potentially can use coal that is economically or technically infeasible to mine because the coal is too thick, deep, dirty, high-ash, and wet; has an excessive angle of dip; or has unpredictable and poor over-burden characteristics that make mining unsafe. Coal can be gasified underground or perhaps liquefied below the earth's surface. Underground coal gasification (UCG) is the prime element of the underground coal conversion program. It enables coal to be converted to gas underground. UCC Coal Resource Recent data show a total U.S. coal resource base of about 6.4 trillion tons. Of this, 1.8 trillion tons might be accessed by UCC, including some minable coal. A commercial UCC process could quadruple roughly the 450 billion tons of U.S. coal reserves, which presently are recoverable by mining. The U.S. coal resources are divided into five sectors. The two most eastern sectors contain relatively thin, swelling bituminous coals that are difficult to gasify. Of the two, the bituminous coal in the Eastern Province is the most difficult because it is deeper than in the Interior Province and the terrain is more rugged. The sector along the Rocky Mountains contains huge amounts of thick low-rank coals that are the most promising for early development by UCC. The coal has good continuity and the terrain is suitable. Low-rank coals such as subbituminous and lignite are highly reactive and tend to shrink and disaggregate when heated. Most of the coal in the Pacific

Jan 9, 1980

By Michael G. Nelson, Hanguang Wu, V. Izaxon, Sukumar Bandopadhyay

The method of mine ventilation involving heated air, which is frequently used in mines in cold regions, does not correspond to the natural conditions prevailing in permafrost regions; it is not economical and leads to a series of complications which can be prevented or greatly reduced by lowering the temperature of the air stream entering the mine until it is equal to or very close to the temperature of the frozen ground This paper presents ventilation design approaches which can be engineered to provide the required micro-climate and thus en- sure the stability bf the ground for placer mines in permafrost. Specifically three ventilation design approaches are presented: 1) design based on minimum thermal impact; 2) design based on positive/negative heat accumulation in the existing mine airways; and 3) design based on heat exchanging devices such as thermosyphons placed above ground.

Jan 1, 1997

By M. Gaft, G. Panczer, R. Reisfeld, L. Nagli

Laser-induced luminescence and breakdown spectroscopy methods appear to be a good tool for real-time evaluation of dolomite content on moving belt. The distinctive feature for time- resolved luminescence is Dy3+ emission in apatite, which is excited by 251 nm laser light and detected with delay time of 5 ms and gate width of 1 ms. The distinctive features for breakdown spectroscopy are 604 nm line of F and P in apatite and 518 nm line of Mg in dolomite. The representative sampling of more than 200 pebbles has been studied by both methods. By luminescence method 100% of dolomite will be identified, while approximately 15-20% of apatite with extremely week Dy3+ luminescence may be mistakenly counted as dolomite. By breakdown spectroscopy the identification is 100% correct both for dolomite and for apatite.

Jan 1, 2002

By Walfrid Been

There are no particular nor unique features of iron mining that are not shared in some degree by all mining ventures everywhere. There are, however, at least three problems that have acquired a special emphasis from a unique cause. That cause is the embarrassment of riches under which the industry is suffering. The most obvious' feature of iron mining today is the copious abundance of usable ore and its almost universal distribution over the face of the earth. By the term "usable ore" is meant ore that is amenable to reduction by the usual and conventional processes. Amidst this plenty, three problems become acutely emphasized: 1. The geographic location that will provide a mine with the best market opportunity. 2. The size of the orebody that will be required to' return the preproduction investment. 3. The geometry and configuration of the orebody that will permit quality control measures to be effectively executed.

Jan 1, 1964

By Robin Adams

After nearly five years of economic hardship, the world copper industry is, in our opinion, about to enter a period of considerable strength. This opinion is based on three main propositions, namely: 1) there are a number of reasons, both general-economic and industry-specific why copper demand is likely to perform significantly, more strongly in the 1980's than in the 1970's; 2) the virtual cessation of investment in new facilities, especially at the mine level, over the past years makes it physically impossible for capacity to expand as fast as demand in the mid-1980's; and 3) there is much less "excess" inventory available than most analysts believe. If we are correct in these assumptions, then the implication is that copper prices will climb up to, and possibly beyond, the cost-justified price level. Combined with general inflation, this suggests a price in excess of $2/pound by 1985 or roughly treble the 1978 average.

Jan 1, 1979

By Janusz S. Laskowski

The most important type of potash ores, sylvinite, contains sylvite, halite and a small percentage of insoluble gangue [sulfates (anhydrite), carbonates (dolomite), and clay minerals]. Primary normal long-chain amines are exlusively utilized to float sylvite from halite, but since mines are also adsorbed by insoluble slimes, the process requires prior desliming. This is carried out either by mechanical means, or by flotation of slimes and blocking off the adsorption centers on clays with the use of polyelectrolytes. Fractional crystallization is employed to produce fertilizer grade KC1 from disseminated ores with a high content of slimes. An extender oil is needed in addition to mine to float coarse sylvite fractions. The process is carried out in a saturated brine at mine concentrations exceeding the solubility limit and is strongly affected by temperature fluctuations. All these important features, which make this process differ from a conventional flotation, result from a high electrolyte concentration.

Jan 1, 1994

By Ross W. Smith, Salim Akhtar

Flotation of oxide minerals, particularly silicates, with cationic collectors differs substantially from classic flotation of sulfide minerals using sulfhydryl collectors. In the case of cationic flotation of these minerals, in general: 1. There is a greater importance of electrostatic interaction between mineral and collector. 2. Longer-chained collectors are required; as a rule at least 10 carbons must be present in the hydrocarbon chain. 3. Moderate concentration of collector is required. 4. Attachment between collector and mineral appears to be less strong even though the system may exhibit a large contact angle. 5. Required conditioning time before flotation is usually relatively short. 6. There is sensitivity to slimes. 7. The collectors usually possess frothing power. 8. Selectivity of cationic collectors for specific minerals is usually poor. Almost any of the oxide and silicate minerals may be floated using cationic collectors. However, specific response to these collect- ors varies with mineral type. The following classes and minerals are

Jan 1, 1976

By P. R. Altringer, R. H. Lien, B. E. Dinsdale

The Bureau of Mines, U.S. Department of the Interior, is investigating biodegradation of cyanide as a means of decommissioning heap leach operations. Biological oxidation decreased weak acid dissociable (WAD) cyanide in process solutions from 50 to 170 ppm down to 0.1 ppm. The tests were conducted both in trickling column reactors, using quartz as the growth surface, and in upflow columns, using activated carbon as the growth surface. In addition, exploratory tests indicated that bacteria will destroy residual cyanide in leached ore. A flowsheet has been proposed for closure of heap leach operations in which the metals processing portion of the plant, possibly the carbon adsorption columns, or a collection pond is used as a bioreactor. The cyanide in the process solution is destroyed in the bioreactor and the treated water is used to rinse residual cyanide from the spent heaps. The water is recycled until the effluent WAD cyanide concentratio� meets discharge standards.

Jan 1, 1992

By Shuji Owada, Keisuke Yasukawa

Electrostatic and/or high tension separation has been used for the separation of rare metal bearing minerals, foods, industrial wastes and so on. Since the separation is achieved by the difference mainly in surface conductivity of particles, the efficiency in humid circumstances is very low, and then, the applicability is considerably limited. We applied the following two-stage pretreatment technique to the separation process, which involves (1) adjustment of surface wettability by adsorbing surfactant on the particle surface and (2) control of the surface conductivity by spraying with electrolyte aqueous solution. The addition of this pretreatment technique has been found to have a great advantage over conventional methods. Separation efficiency is achieved even in humid circumstances and selectivity can also be obtained between insulator minerals. The selectivity is very similar to that of flotation in wet process. This separation technique will be applied not only to mineral processing but the processing of industrial waste such as electric and electronic waste materials.

Jan 1, 1995

By R. A. Metz, Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (1), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profitability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus minimum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12). Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-lowgrade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1992

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

The purpose of this section is to demonstrate the development of a method that we believe can be used for simulating and predicting mine water inflow. It will also show how important it is for personnel who model mine water inflow to have considerable experience with this subject, and with the conversion of hydrogeologic frameworks to conceptual models. Several of the models described previously are said to be three dimensional. However, closer examination reveals that there is considerable limitation in the variation possible in the third dimension. A true three-dimensional analysis would require a very large program. Therefore, it seems more reasonable to develop a procedure by which information from a two-dimensional model can be used for analysis of a three-dimensional problem. For a vertical mine shaft this is relatively easy because most two-dimensional models have an option for axisymmetric flow. If it can be assumed that axisymmetry exists about the centerline of the shaft it can be modeled easily as has been illustrated previously. For a horizontal drift or tunnel the analysis is not nearly as straightforward. However, the flow can be divided into two parts as indicated in Fig. 59. Part 1 considers the flow into the side of the drift while part 2 considers the flow into the end of the drift. To calculate the flow rate into the entire drift from the unsteady flow into a unit length the instantaneous flow per unit length may be integrated over the total length of the shaft. This procedure is shown in Fig. 60. The radial flow through any incremental length of the shaft is q. The total flow rate, Q, into the drift at any particular time is given by: The discharge q is a function of the time during which a particular segment has been draining that is dependent on the location x. In other words, a location at x = 0 has been draining during the entire time of expansion while a segment at x = L has only started to drain. Therefore, the time of draining at any point, x, is given by: [ ] where t' is the time over which the mine has been expanded. To determine the relationship between q and t it is necessary to refer to Polubarinova- Kochina (1962, pp. 566-569) and the equation given for the unsteady flow problem. This is: [ ] where q is the steady flow, q,, q, are the coefficients, and t is the time since drainage began. The expressions for q0, q1, and q2 are given as:

Jan 1, 1986

By M. Zhang, H. Ni, F. Zheng, M. Guo

"In the present work, we investigate the effects of roasting temperature and time on the transformation of the titanium (Ti)-bearing phase and magnesium aluminate spinel (MgAl2O4) phase in Ti-bearing electric arc furnace molten slag during the molten sodium hydroxide (NaOH) treatment process. With applied temperature of 700 °C for 60 min, and a mass ratio of Ti-bearing electric arc furnace slag to NaOH, or Rslag/NaOH, of 1:1.2, the conversion ratios of the Ti-bearing phase and the MgAl2O4 phase reached about 98.3 percent and 88.1 percent, respectively. Moreover, investigation of the reaction kinetics indicate that the conversion of these phases during the molten NaOH treatment process is a typical solid-liquid reaction, and it can be theoretically described by the unreacted shrinking core model. The conversion ratio of the Ti-bearing phase was found to be controlled by the chemical reaction step, with an apparent activation energy of 41.76 kJ/mol. The decomposition of the MgAl2O4 phase was found to be controlled by the surface chemical reaction in the early stage of 0 to 20 min from start of reaction, changing to diffusion control in the later stage of 20 to 60 min. The apparent activation energies of the process were determined as 55.88 and 22.72 kJ/mol, respectively.IntroductionChina has rich mineral resources of vanadium–titanium magnetite ore deposits, and the iron and steel industry in the country produces millions of tons of titanium (Ti)-bearing slag every year. The problem of how to dispose of the Ti-bearing slag has attracted much attention, and various approaches have been developed, including the acid process (Li et al., 2007; Zhang et al., 2011; Zhang and Nicol, 2010; Li et al., 2011), alkali fusion process (Han et al., 2012; Middlemas, Fang and Fan, 2013; Nayl, Awwad and Aly, 2009; Lasheen, 2008; Zhang, Qi and Zhang, 2009), carbonization and chlorination process (Liu et al., 2008; Tao et al., 2012; Fu et al., 2012; Liu, 2009), selective crystallization process (Lei and Xue, 2008; Li et al., 2011; Zhang et al., 2007a, 2007b). However, these processes have been beset by various functional, economical and ecological problems (Yu and Chen, 2010). With environmental protection regulations becoming increasingly rigorous, and demand growing for the comprehensive utilization of resources, the focus is on methods to treat Ti-bearing slag with high efficiency and mild reaction conditions (Nazari and Riahi, 2011; Wu et al., 2013)."

Jan 1, 2017

By David A. Holmes

Zeolites are crystalline, hydrated aluminosilicates of the alkali and alkaline-earth metals. Their crystalline framework is based on a three dimensional network of SiO4 tetrahedra with all four oxygens shared by adjacent tetrahedra. The alkali and alkaline-earth cations are loosely bound within this framework structure and may be exchanged, to varying degrees, with other cations or molecular water. Most zeolites can be reversibly dehydrated. At present, close to 40 natural zeolite minerals have been identified and over 100 zeolites have been synthesized. Zeolites are a family of hydrous aluminosilicate minerals whose physical properties provide unique features for a wide variety of practical applications. Zeolites have been known for more than 200 years, but it was not until the middle of this century that their economic potential and widespread occurrence in sedimentary rocks were known. Since 1950, literally thousands of articles and patents have appeared in technical literature, and commercial pro¬duction of natural zeolites may now exceed 1 Mt worldwide an¬nually. In the past 40 years, the zeolite mineral group has changed from that of a museum curiosity to one of an established industrial mineral commodity. According to Breck (1974), zeolites are characterized by the following properties: 1) High degree of hydration, 2) Low density and large void volume when dehydrated, 3) Stability of the crystal structure of many zeolites when dehydrated, 4) Cation exchange properties, 5) Uniform molecular-sized channels in the dehydrated crys¬tals, 6) Ability to adsorb gases and vapors, 7) Catalytic properties. All zeolites are molecular sieves, but not all molecular sieves are zeolites. Activated carbon, activated clays, alumina powder, and silica gels are also molecular sieves. Molecular sieves are materials that can selectively adsorb molecules on the basis of their size, shape, or electrical charge (Clifton, 1987). Commercial application of zeolites is based on the following physical properties: molecular sieving, ion exchange, adsorption, and catalysis. Potential commercial utilization includes hundreds of possible applications. The principal uses at this time, however, are as fol¬lows: -Ammonium-ion removal in waste stream treatment, sewage treat¬ment, pet litter, and aquaculture, -Odor control, -Heavy metal ion removal from nuclear, mine and industrial wastes, -Agricultural use, such as soil conditioner and animal feed supplement. An estimated 800 kt to 1 Mt of natural zeolites were produced and consumed worldwide in 1991. Eight small operators reported collective production and consumption of an estimated 18 kt of natural zeolites in the United States during 1991, mostly for feed additives and for odor control, aquacultural, agricultural, catalytic, and ion exchange applications (Eyde, 1992).

Jan 1, 1994

By Bernard L. Merlo

INTRODUCTION Modern mine management is constantly striving to develop improved mining methods which will not only increase productivity, but will also provide a safer work environment and make the most efficient use of our natural resources. Progress can be seen in this area by tracing mining methods through history. Work that was originally done with hand tools progressed with the age of mechanization to conventional mining, matured to continuous and longwall mining, and the advent of short¬wall mining, a combination of continuous and longwall, has recently been seen. This new method of shortwall mining provides a promising method for safe and ex¬pedient coal extraction. This study focuses upon the planning, design de¬velopment, and operation of the first producing short¬wall unit in the US which was operated in the Pittsburgh coal seam in northern West Virginia. MINE DESCRIPTION The mine is located in northern West Virginia at Grant Town in the heart of the high volatile Pittsburgh coal seam. Originally opened in 1901, the mine has be¬come one of the largest in the US, encompassing an area of approximately 103.6 sq km (40 sq miles). The seam being mined was the high volatile Pittsburgh seam which ranged in thickness from 1.8 to 2.96 m (6 to 91/2 ft). Three types of mining methods were employed: con¬tinuous mining, longwalling, and shortwalling, with each type accounting for 50%, 30%, and 20% of total pro¬duction, respectively. A total of 630 contract and super¬visory personnel were employed to produce the mine's designed annual tonnage of 1.8 Mt (2 million st). There were ten producing sections, eight continuous miners, one longwall, and one shortwall, operating a total of 22 unit shifts per day. Eight shafts, varying in depth from 80.2 to 194.2 m (263 to 637 ft) provided entrance to the Pittsburgh coal seam. These shafts were used for ventilation, portaling, coal hoisting, supply, and emergency escape hoisting. Ventilation was achieved by means of three exhaust fans which provided 707.9 m3/s (1.5 million cfm) of air to the mine and removed approximately 141 584 m3 (5 mil¬lion cu ft) of methane per 24 hr period. The room-and-pillar method was used for secondary recovery with designed barrier blocks between panels of 61 m (200 ft) in width. A double row of blocks was left around each gob area to provide efficient bleeder action

Jan 1, 1982

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 J. O. Marsden

The 1980s saw a revolution in gold extraction technology, paralleled only by the implementation of the cyanide leaching process in the late 1800s. This revolution was led by the widespread adoption of carbon adsorption processes for gold recovery, heap leaching of low-grade ores and development of a variety of oxidation processes to treat refractory ores. In the late 1980s, gold production in almost all regions of the world increased, with the notable exception of South Africa. This geographical variety is creating a greater diversity of challenges for gold metallurgists. This is due to a wider range of ore types, new and varied environmental regulations and legislation, and a requirement to select the most appropriate technology to suit local needs. North America has led this technological revolution in the gold industry and is responsible for commercialization of most of the major innovations: carbon-in-pulp, heap leaching, whole ore pressure oxidation and whole ore roasting. Process selection trends A general trend in precious metals extraction technology has been the treatment of ores with progressively lower grades. The lowest cost treatment, dump leaching of material at run-of-mine size, has been increasingly applied. ZortmanLandusky, in Montana, is successfully leaching material above a cutoff grade of 0.2 g/t (0.006 oz/st) at an average rate of about 28 kt/d (31,000 stpd). At this scale of operation and in the absence of any cyanide-consuming species in the ore, process operating costs for a ge¬neric run-of-mine leaching operation are estimated to be between 60 and 75 cents/ t (66 and 77 cents/st). One of the largest heap leach opera¬tions treating crushed ore is the Round Mountain Mine in Nevada. It processes about 45 kt/d (49,000 stpd) at an average grade of 1.1 g/t (0.032 oz/st) at a nomi¬nal crush size of 80% -19 mm (-0.7 in.). Round Mountain also treats lower grade ore above a 0.2 g/t (0.006 oz/st) cutoff by dump leaching at run-of-mine size.

Jan 1, 1993