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By Richard R. Klimpel

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

Jan 1, 1988

By D. A. Norrgran, J. A. Marin

Introduction The recent development of rare earth permanent magnets has revolutionized the field of magnetic separation. The advent of rare earth permanent magnets in the 1980s provided a magnetic product that was an order of magnitude stronger than that conventional ferrite magnets. This allowed for the design of high-intensity magnetic circuits that operated energy free and that surpassed electromagnets in the strength and effectiveness. New applications and design concepts that focused on the mineral and metal processing industries have evolved. This technology led to the development of various magnetic separators specifically designed for mineral processing applications. Applications that were not previously considered are now being used in primary mineral upgrading, recycling and secondary recovery. Historical perspective Lodestone was the first naturally occurring permanent magnetic material known. Lodstone was most likely used to upgrade iron ore by early civilizations. By the 1600s, the early magnet technology had advanced to quench-hardened iron-carbon alloys. The practical significance of magnetic separation was formally recognized in 1792 when an English patent was issued for separating iron ore by magnetic attraction. By today's standards, carbon steel is a very poor magnet material. It is easily demagnetized and has a very low energy product of much less than I MGOe (Million-Gauss-Oerstads). This was state-of-the-art technology for almost 300 years until chromium was added to magnet feedstock, which resulted in a three-fold increase in the energy product. The well documented addition of cobalt to permanent magnets in 1917 initiated the 30-year era of "Alnico" magnets that at the time provided a superior magnetic energy product. Since then, the science of magnetism has advanced rapidly and is now considered a highly developed branch of physics and material science. Permanent magnets have had an extremely long history. Figure 1 presents a chronology of permanent magnets that illustrates the increase in energy product. Amazing developments in material science have taken place in the last two decades. The gradual advancement of permanent magnet technology was shattered in 1967 with the initial development of samarium-cobalt (rare earth) magnets. Since that time, the advent of neodymium-boron-iron magnets provided such an increase in energy product that new design concepts were considered. New avenues of study were introduced by the complexities in the material science and physics involved in describing these new permanent magnets. Furthermore, applications for permanent magnets that were previously not considered were now viable. Rare earth elements Rare earth elements have claimed the attention of scientists for the past century. These elements were originally termed "rare" because they were thought to be quite scarce. Since then, however, geological studies have shown them to be relatively abundant. The discovery and identification of rare earth elements is complicated by the inherent difficulties in separating them from each other. The rare earth elements comprise the fifteen transition elements of Group IIIB, Period 6, of the periodic table. These elements extend from lanthanum to lutetium and are commonly called the lanthanide series. Samarium and neodymium are the two most common elements used in the commercial manufacture of rare earth permanent magnets. Commercial grade rare earth magnets There are only a few common types of rare earth magnets that are considered in the circuit design for magnetic separators. Early rare earth magnets of commercial significance (introduced in 1970) consisted of the first generation of sintered SmCo5. The energy product of these magnets ranged up to 23 MGOe, which provided the initial impetus to the field of high-energy permanent magnets. Although these magnets did not produce the extremely high magnetic field strengths of current rare earth magnets, they were relatively temperature stable. Containing 66% Co, they are the most expensive of the basic commercial rare earth permanent magnets. Their use is limited today because they are being replaced by second and third generation rare earth permanent magnets.

Jan 1, 1995

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thor¬ough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield reve¬nues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equip¬ment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Ar¬ranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1989

By G. Kemp Williams, Peter H. Dohms, Barry A. Hillman, Laurie A. Williams

Introduction In 1969, the US devalued the dollar. This action, over the ensuing four years, resulted in the deregulation of the price of gold. It also resulted in the rebirth of the US gold mining industry. The extreme inflation of the 1970s led to a dramatic increase in the price of precious metals. Consequently an increased interest in the exploration for and mining of gold has occurred. The enthusiasm began among the few existing producers who had survived on marginal profits during a prolonged period of fixed prices and constantly rising costs. Then came the newcomers. They saw the opportunity to enter an industry that historically had been characterized by small operations with low capital costs. It was not until the late 1970s that large mining companies recognized that investment in gold projects could generate cash flows that previously were associated only with world class deposits. Today, gold mining has tremendous appeal for the entire mining community, despite the dramatic price drop from 1980-81 levels. There are even major firms with no mining experience attempting to gain a foothold. But how easy is it to enter the US gold mining industry? What are the basic factors that influence the economics, and what are they likely to be for the remainder of the decade? How has the industry changed, and what is the composition of the active companies? An examination of the industry will help to answer these questions. US Gold Mining History The current gold rush is actually the fifth in the 208-year history of the US. The first recorded commercial mining of gold occurred in the Southeast in 1799. It produced from small, yet relatively rich placer deposits. The famous California gold rush of 1848 placed the US as a major producer with 62.2 t (2 million oz) of output by 1850. More than 93.3 t (3 million oz) was mined by 1853, equivalent to 63% of world production. But the richest placers were soon depleted. By the end of the Civil War, annual production had declined to well below 62.2 t (2 million oz). The government then relaxed prices and the industry again began to recover. The 1870-80s saw the great mines of the California Mother Lode and Grass Valley brought into production. This was followed by development of Nevada's Comstock, which later became one of the world's largest producers. Discovery of gold at Cripple Creek in 1892, coupled with technological advances in the use of large dredges in the California placers, resulted in significant production increases. And, with the expansion of the Homestake mine at Lead, SD, in 1898, US output again surpassed the 93.3-t (3 million-oz) level. By 1905, produc¬tion climbed to 124.4 t (4 million oz) with the addition of the Alaskan deposits, new output at Goldfields, NV, and the introduction of the cyanidation process. During World War I, the government forced the shutdown of gold mines. By 1920, total production had dropped to little more than 62.2 t (2 million oz). Many mines remained closed until the Great Depression. This, combined with the devaluation of the US dollar in 1934, resulted in a gold price increase to about $1.13/g ($35 per oz). Then began the fourth gold rush. By 1937, output had again surpassed 124.4 t (4 million oz). This was due to an increase in primary production and new byproduct production from base metal operations. Unlike the previous period of growth, however, this period was short-lived. The imposition of Executive Order L-208 in 1942 once again forced the shutdown of most mines. At the end of World War II, cumulative US gold production had

Jan 12, 1984

By Spencer R. Titley

"For his pioneering contributions to the understanding of the geology of porphyry copper systems" Recipient Of The 1997 D. C. Jackling Award SPENCER R. TITLEY Introduction The subject of this lecture, a late century perspective on porphyry copper geology, is one of current interest and importance as the search for these ores remains intense. It is presented in the belief that an appreciation of our present understanding of the subject requires it to be viewed historically, thus providing a basis for evaluating appropriate directions for continued study. At the outset, it is important to note that the technology of mining and extractive metallurgy has developed ore from these low grade orebodies, starting with D. C. Jackling, and has continued to create them as copper grade and price have diminished. But the continued search for them is a geological problem that responds to growing expectations of people and societies; exploration based upon ever increasing knowledge of the habits of the ores has been very successful. This is the story of evolution of geological thought and practice in porphyry ore search across an interval of time beginning with the bulk mining of "low grade" copper ores at Bingham and ending with late Century concepts. "How did we get here?" The lecture concerns the ideas of many different people. The ideas and events discussed reflect a bias on North American thought, where the great low grade orebodies were first recognized as such and where many of the basic geological ideas were developed. A very great number of geologists have contributed to this understanding. In no sense is it meant to minimize the contributions to the story that have evolved in the study and mining of ores of the Andes, whose development and expansion has been nearly synchronous with that in North America - nor to minimized comparable contributions made in the late century from studies of the newer discoveries of deposits in other regions, especially the southwestern Pacific regions. In expansion of this topic, I have found that many important concepts are marked by appropriate publication but some are not, or I have not been able to discover documentation. Many notions appear to have become conventional knowledge without publication, mostly being geological viewpoints developed synchronously from similar observation. Thus, apologies are extended in advance to those whose work I have might have missed and to those unknown to me whose unpublished ideas may have seeped into the pool of common knowledge without appropriate recognition. The analysis here addresses events of two time segments, an older one extending from the developments at Bingham in the early 1900s through the Second World War, ending about 1950: the second period extends from that time to the present. This separation is seen in Figure 1, which traces numbers of publications through the period as extracted from some 700 publications, again with a North American bias, from an annotated bibliography of porphyry copper geology in preparation. During the early period, exploration was "immature" in North America, that is, geologists looked for new districts and sites of the deposits. Outcrops were sought that carried the signatures of former mineralization and regional controls related to structural geology were closely studied. In the later period, North American exploration had "matured" into a study of districts, rather than a search for them. As the studies of these ores intensified during the younger period concommitant with increasing demands for the metal, they did so with a broadened understanding of some habits of occurrence, of hydrothermal alteration, mineral zoning, petrology, and geochemical and geophysical signatures. Outside of North America, search remained "immature" but closely related to the

Jan 1, 1998

By David W. Cannon, Russell V. Mann

In a dispersion of particles in liquid a net charge will develop at the particle-liquid interface. This surface charge is usually due to the adsorption of charged material from solution. The existence of this surface charge gives rise to the formation of the electric double layer of counter-ions which surrond the charged particle. The particle surface charge and the electrostatic repulsion which exists between similarly charged particles is the primary stabilization mechanism for lyophobic colloids (1). The separation of charge which occurs at the particle-liquid interface gives rise to several dynamic phenomena associated with colloidal systems or with solid-liquid interfaces in general. These phenomena are known as electrokinetic phenomena and the four classic electrokinetic phenomena are; electrophoresis, electroosmosis, streaming potential, and sedimentation or Dorn potential. The actual driving force for electrokinetic phenomena is not the surface charge per se, but the charge at the interface between the liquid which is hydrodynamically bound to the particle surface and the bulk fluid. This interface is known as the slipping plane or the plane of shear and the potential at this interface is the zeta potential (2). The factors linking the electrokinetic phenomena is that they involve a relative motion between the liquid and the charged particle or solid surface and the driving force is the zeta potential of the solid. In addition to the four classic electrokinetic phenomena there are two additional electrokinetic effects in disperse systems; the electro-acoustic effects. When an alternating electric field is applied to a cooloidal dispersion the particles will move in the field due to their net zeta potential. If there is a density difference between the particles and the fluid this motion will result in the development of an acoustic wave. The effect was discovered at Matec and has been termed the Electrokinetic Sonic Amplitude or ESA (3). ESA is the pressure amplitude generated by the colloid per unit applied electric field strength and has SI units of pascals per volt per meter. When an alternating pressure field (acoustic wave) is applied to a colloidal dispersion the inverse of the ESA effects occurs. A density difference between the disperse phase and the continuous phase leads to a relative motion between the particles and the surrounding liquid. This means that there will be a periodic displacement between the charged particle and the oppositely charged counter-ions in the electric double year. This displacement results in the development of an alternating dipole moment at the frequency of the applied field. This effect is termed the Ultrasonic Vibration Potential or UVP and was first predicted for electrolyte solutions by Debye in 1933 (4). UVP is measured in units of volts per unit velocity amplitude of the applied acoustic field or volts per meter per second. In 1938, Rutgers (5) and Hermans (6) pointed out that the effect would also be present in colloidal dispersions. A detailed theory for UVP effects in colloids, also called CVP, was first presented by Enderby in 1951 (7). Extensive studies of the UVP in electrolytes have been carried out by Yeager et al (8). Recently, O'Brien (9) has developed a general theoretical treatment of electro- acoustic effects in colloids and has derived a reciprocal relation linking ESA and CVP effects. The most commonly studied electrokinetic phenomena is electrophoresis. Electrophoresis is the movement of charges particles in an applied electric field. The velocity of the particle divided by the applied electric field strength is the electrophoretic mobility of the particle. The zeta potential can be calculated from this mobility (2). The magnitude of both the ESA and CVP effects are directly proportional to the electrophoretic mobility of the particles. The mobility determined by the two electro-acoustic effects is the dynamic or AC mobility of the particles.

Jan 1, 1990

By D. Nilsson, B. Aaro

To find out if there are any potential savings in "in-the-wall haulage for open pits," the Swedish companies ASEA and Kiruna-Truck in 1984 gave us financial support to study this solution in more detail. In the study, the use of in-the-wall haulage was studied for three hypothetical open pits with different sizes and shapes. Annual ore production rates also varied. A part of this thorough study was published by the Swedish Mining Research Foundation (see Report F 8444, "The Use of Electric Trucks and In-the-Wall Haulage in Open-Pit Mining," in Swedish with an English Summary). Although in-the-wall haulage is of interest in some open pits, the local terrain is important for profitability. The authors do not think in-the-wall haulage is of any major interest for the mining industry. The following is a summary of some of the author's findings. To place the haul road in the wall is not of interest in open pits with declining metal contents in the bedrock. In such mines, the volumes from the haul roads will not yield revenues. The cost per m3 for an underground haul road is much higher than for a haul road in the pit. This means that it is only of interest to place the haul road in the wall when mining gets deeper than about 50 m (165 ft). In-the-wall haul roads will reduce flexibility in the pit, and it will make necessary the use of smaller equipment with lower productivity and higher costs per ton. As Hustrulid, et al. show (Table 4), excavation savings of material hauled are very low, $0.10 to $0.25/t (0.09 to 0.27 per st). The extra operating cost, due to lower productive equipment, will normally be much higher and thus destroy the whole idea with in-the-wall haulage. If electric trucks are profitable for in-the-wall haulage, it is normally also profitable to use electric trolley assist for trucks on the haul road in the pit. But the profitability of using electric power is different in different countries, and depends on the relation between the cost for electric power and diesel fuel. In the US, diesel fuel is inexpensive compared with electricity, but in a country like Sweden, diesel fuel is much more expensive. In most open pits, the trucks have to move a considerable distange from the loader until they reach a final haul road in the wall. A trolley line along such temporary haul roads will be exposed to flyrock. It is normally less expensive to perform rock support from the open pit than from underground ramps in the wall. In Fig. 8, Hustrulid et al. gives the impression that the underground haul roads will be very close to each other. This is seldom the case. Figure 1 shows a haul road in the wall. The haul road passes through each cross-section only once. Arranging a reliable dewatering and rock support system from only one underground ramp is probably impossible. Many more drifts are therefore necessary. In our report, we also studied what would happen if the final pit slope is increased by 5°, using underground drifts in the wall. Our conclusion was that, the extra cost for drifts, rock support, pumps, etc. destroys the whole idea, and that it was better to accept a higher stripping ratio with the haul road in the pit and to use conventional low-cost open pit equipment. Finally, we think it is would be difficult to try to reduce the safety factor when determining the slope angle by moving the haul road in the wall. Minimizing the risk for pit wall collapse is also important with in-the-wall haulage, primarily because men and equipment will be working in the pit, but also to guarantee many accesses between the pit and the haul road in the wall.

Jan 1, 1990

By G. Sciocchetti, G. F. Clemente, F. Scacco

INTRODUCTION Radon, thoron and their airborne daughter products may result in a significant biological hazard when they are concentrated in enclosed places such as mines, houses and caves. High concentrations of radon and its particulate daughter have long, been known to be a causative agent for lung cancer to uranium miner: (Archer 1973). More recently increasing concern has been expressed on the possible health hazards associated with long term exposures over large population groups to low levels of radon daughters (Cohen 1980, Stranden 1980). Due to its peculiar geological features, Italy has various natural environments where the content of uranium and thorium in the soil may increase underground radioactivity. The presence in Italy of many underground waters rich in radon may be considered a further important source of radon in underground mines. A preliminary survey of the radon and radon daughter air concentrations in italian underground mines was started in 1971 (Bottino 1973). Furthermore many studies (Sciocchetti 1978, Sciocchetti 1980, Clemente 1980) have been performed during the last ten year to characterize the natural radiation environment in Italy. It has been also demonstrated that 21OPb/21OPo skeletal burden in spa workers and non-uranium miners can be heavily increased by the exposure (short or/and prolonged) to high levels of radon daughter air concentration (Clemente 1980). All these studies have been of great help to put in the right prospective the radiation hazards linked to the underground working activity in non-uranium italian mines. Two main actions have been prosecuted from 1971 until today in order to evaluate in greater detail such radiation hazards and then, when needed, to obtain their reduction to levels which may be considered acceptable. The two actions may schematically described as the following: - Measurement of the radon and daughter (and thoron and daughter when needed) concentrations in most of the italian non-uranium mines. The assessment of the radiation hazards linked to both the annual average lung exposure and the 210 Pb - 210 Po annual average dose to the skeleton is then performed for each considered mine by taking into account the main parameters characteristic of the mine. - Set up of a national regulatory recommendation on the permissible limits for the radiation exposure of workers due to underground mining activity. Such recommendation have been established in 1978 and have been designed in order to implement a national radiation protection practice during the underground mining activity. In this paper a summary description of the results obtained during the execution of these two actions will be given. MATERIAL AND METHODS Measurement of Radon, Thoron and their Daughter Products The excution of the planned survey requires the availability of specific techniques and instruments for measuring the different parameters characterizing the radon, thoron and their daughter exposure in mines. A reference laboratory was set up for the development and the standardization of the measurement procedures. Measurements of radon, thoron and daughter concentrations are performed by various methods: instanteneous or grab sampling methods and continuous read out monitors. The radon concentration is usually measured by a modified version of the two filter method. Mine air is drawn into a polyethylene bag and the sample is there after injected into a two filter tube; after sampling the exit filter is counted by the alpha spectrometric technique. The measurements is completed in a total time of less than 25 minutes at a sensitivity level as low as 3.7 Bq/m3 for a tube volume of 15 1. The measurement of radon and thoron daughter concentration and of the relevant potential alpha energy concentration is performed by means of a portable spectrometric apparatus which was developed at our laboratory (Sciocchetti 1981). The device allows multiple operating modes: instantaneous or brief sampling period and automatic cycles; computer data processing is then provided. The instruments that measure continuously radon concentration over long periods of time are of two types: a continuous scintillation monitor and an automated two filter

Jan 1, 1981

By H. V. S. Tingley, D. Helfant Ghose

"Ten years ago," said a mining industry personnel manager, "we used to wine and dine new mining engineering graduates. Now I can pick and choose among men with 10 years of experience." Comments like that are common throughout the industry today. During the 1970s, the US and Canada did indeed suffer from a severe shortage of mining engineers, especially experienced ones. Moreover, it seemed likely this North American shortage would worsen. Increased production of minerals and coal was pushing demand for mining engineers upwards, and there appeared to be substantial new employment prospects for them in the nascent synfuels industry. The oil majors entering the mining industry were pushing for an increased professionalization of mining management, heralding a trend towards higher intensity in the use of mining engineers. In response to the shortage, mining schools expanded their programs dramatically. By 1982, American schools were churning out more than 600 new mining engineers per year, compared to barely 200 in 1970 (Table 1). The timing of the explosion in the number of mining engineers could not have been worse. The downturn in the economy and the resulting slump in the mining industry coincided almost exactly with peak output of mining engineers. The market for mining engineers shifted from shortage to surplus with startling rapidity. The new crop of engineers, no longer courted by the mining companies, struggled to find any mining jobs. At the same time, experienced mining engineers found themselves thrown into the job market, some for the first time in decades. The current oversupply of mining engineers is assumed to be a temporary phenomenon. As the economy recovers, the demand for mining engineers is expected to pick up at least enough to absorb today's surplus. Not so, according to a recent study by Fenvessy and Schwab, a general management consulting and executive recruiting firm. In 1979, Fenvessy and Schwab was commissioned by a group of major mining companies to conduct a worldwide study of supply and demand for mining engineers to the year 2000*. The study began at the height of the North American shortage. One major objective of the study was "to identify means by which business and educational institutions may cope with and alleviate a future shortage of mining engineering talent." The findings, however, ran directly counter to expectations. The current oversupply will not disappear with economic recovery, the study concluded. In fact, the US and Canada are facing a massive oversupply that will probably last until the end of the century, unless enrollments in schools are reduced. Study Method Fenvessy and Schwab's conclusions are based on supply and demand models for mining engineers that it developed in the course of the study. Specifically, predictions for the supply of mining engineers are based on a survey of the 54 mining schools that account for about 90% of the world's annual output of mining engineering graduates. This survey asked for data on historical and projected number of graduates, obstacles to further expansion, changing trends in the use of mining engineers, and graduate career patterns. Fenvessy and Schwab developed three mining engineering supply scenarios from the survey data. The high case, considered unlikely for most of the industrialized world, is based on a continuation of the 1970's soaring growth rates. The middle case was established from the mining schools' own estimates of future growth in the number of graduating mining engineers. Finally, a "low growth" case for the US and Canada assumed that the number of graduating engineers will remain at current levels. Mining engineering demand was forecast using a complex model based on the demand for 50 major minerals. Formulas were

Jan 9, 1983

By R. F. Holub, E. Franco, T. B. Borak, K. J. Schiager

INTRODUCTION According to a recent analysis by Schiager et al.(1981), inhalation of radon daughters continues to contribute the largest risk from exposure to radioactive materials in underground uranium mines. The American National Standards Institute (1973) has published a concensus standard for radiation protection in uranium mines. According to this standard, a monitoring system for radon daughters must be capable of measuring the annual accumulated exposure in working level months within an uncertainty interval of 50% at the 95% confidence level. To satisfy this criterion, uranium mining companies in the USA have adopted a procedure consisting of single or periodic measurements of the air concentration of radon daughters. All of the measurement methods sample a known volume of air through a filter. After a specified delay interval, the filter is analyzed for radioactivity. The exposure rate from total airborne potential alpha energy in units of working level (WL) is estimated from the measured activity on the filter. For control purposes, this measurement establishes whether or not a work area is safe for occupancy by miners. For routine dosimetry, the concentration estimated from the grab sample is assigned to the appropriate work area. It is then assumed that the concentration of radon progeny is equal to the average of the two measurements at the beginning and end of the specified interval. Meanwhile, the work locations and exposure times of each miner are recorded on the daily time card and verified by the shift supervisor. Most cards provide space for 15-minute time intervals. The exposure in working-level hours (WL-hr) for each shift is obtained by summing the time spent in a given location multiplied by the exposure rate (WL) assigned to that location.The exposure of an individual miner to radon progeny is summarized on a weekly basis and commonly reviewed each month. These values are then accumulated in units of working level- months (WLM) during the calendar year. EVALUATION TECHNIQUE AND RESULTS The evaluation of grab sampling methods for monitoring exposure to radon progeny is presented in two sections. The first section examines each method individually according to accuracy, precision, and uncertainties attributed to human errors. The second section includes an analysis of procedures which are common to all measurement methods. The ultimate goal is to determine whether a specific measurement method, or grab sampling in general, can fulfill the criterion that the annual accumulated exposure of each miner be known to within ± 50% at the 95% confidence level. In addition it is essential that the above criterion be satisfied at an acceptable cost to the mining company. The results of each testing procedure are presented in terms of the contribution of the relative uncertainty in the estimates of cumulative WLM. Each of these are then combined to form a figure of merit for the complete monitoring system. The method of combining statistical uncertainties and interpretation of the results is subject to some debate. However, since all of the tested parameters enter into the final computation of WLM multiplicatively, we have used the commonly accepted procedure of combining the squares of the relative uncertainties. We have also assumed that each of the uncertainties is derived from independent random variables and the covariance of any two variables is zero. A description of each grab sampling measuring method is given in Table 1. Also included is the equation used to compute WL from the measured quantities. Accuracy (U A) Our definition of accuracy is based on the uncertainty inherent in any method which attempts to estimate potential alpha energy without unambiguously determining the concentration of the three unknown radioisotopes, RaA, RaB and RaC. (For simplicity, the RaC and RaC' nuclides are referred to in combination as RaC.) The inherent uncertainty strongly depends on the mixture of these three components at the time of measurement.

Jan 1, 1981

By Joseph R. Wojcik

Introduction Placer deposits account for one-half to two-thirds of the total gold produced worldwide. As many as 311 t (10 million oz) have been produced from some major districts. Several others are credited with 155 t (5 million oz). All the major districts and many lesser ones have similar characteristics. Consideration of these characteristics offers insight into the origin of gold in the deposit. Knowing the gold deposit's geologic history leads to useful conclusions in exploring for placer deposits and for deposits in consolidated areas. Placer Deposit Characteristics Table 1 lists 17 placer gold districts around the world and their characteristics. These include type and age of, mineralized source rocks, associated intrusive rocks, and age of intermediate sedimentary cycles. Although quartzite or sandstone occurs in the bedrock sequence in four of the 17 districts, slates, phyllites, or schists are present in all of the districts. Additionally, in all but three, the slates are pre-Mesozoic. And, in six of the remaining 14, the slates are Precambrian. Rocks intrusive into the slates are all of intermediate to acidic composition and are emplaced as stocks, dikes, or sills, rather than being batholithic in proportions. Intermediate interceptors have not been recognized in only three of the 17 districts. Data from China and Mongolia are inconclusive as to this characteristic. Black Shale Gold Higher than average gold values are found in modern sedimentary rocks in coarse granulites, conglomerates, sandstones, and gray-wackes. Not so predictably, higher than average gold values also occur in black shales and carbonaceous argillities. Li and Shokena (1974) found that, in the Proterozoic sedimentary and equivalent metamorphic rocks of the Enisei Ridge, carbonaceous and graphitic phyllites and schists were particularly enriched in gold. Korotayeva and Polikarpochkin (1969) reported gold content in organic rocks to be enriched up to 15 times more than the organic free shales and siltstones of the Nerchinsk Zavod region of eastern Transbaikal. Black shales can be gold interceptors as are rocks of the granulite facies. Gold occurs in shales in particulate, free form, in pyrite, adsorbed on clays, and precipitated on carbon. In copper-bearing shales of Poland's Fore-Sudetic monocline, Kucha (1973) reported gold present in organic compounds as well as in silver bearing minerals. Anoshin (1969) found a strong correlation between gold and the pelitic fraction of sediments in certain rocks of the Atlantic Ocean basin. Heavy metal absorption on clays results in agglomeration to the extent that the particles settle with the heavy mineral fraction in enriched strata. Piper and Graef (1974) think that gold in sediments on the flank of the East Pacific rise is of lithogenic origin. Crockett (1973), however, proposed a volcanogenic exhalative origin. A strong correlation between gold and volcanic activity exists in rocks of the Canadian Shield where tuffaceous sediments and porphyries are more frequently enriched in gold than rocks with no volcanic components. Volcanic rocks of intermediate to acidic composition are considered here to be the primary source of gold in black shales. Gold is concentrated in shales as detrital free grains with diameters less than 50 µm (270 mesh), as precipitates, and in pyrite or other sulfide minerals. This is the first step in the history of gold placer deposits. Grain Size Modification Gold, as free grains with diameters less than 50 µm (270 mesh), is too light to be concentrated hydraulically into placer deposits. Some process has increased the average size of the gold grains. Mobilization during regional metamorphism converts most of the gold into particulate grains or entrains it in microfractures in the sulfide minerals. With moderate folding, small, white, vitreous quartz veins develop. As deformation continues, larger, more conspicuous quartz veins are formed. The larger veins are folded and faulted and contain native gold as minute specks and flakes not visi-

Jan 11, 1984

Disposal of solid and liquid wastes from mining and milling operations has become a focus of attention in recent years. Consequently, the geotechnical engineer is more frequently associated with design of disposal facilities. Jack Wulff, a representative of one consulting firm with experience in this field, talks with ME about the geotechnical engineer's contribution. Mr. Wulff, mine and mill waste disposal today is receiving more attention-and many more dollars-than in the past. Regulatory agencies have promulgated demanding requirements concerning mill waste (especially uranium mill waste) and specialist consultants are frequently being called in to design disposal systems. What can the experienced geotechnical engineer add to the development of waste disposal facilities? It's true that geotechnical engineers have recently become more visible in the minerals sector, but we've been around mining for a long time-as foundation engineers for mills and mine support buildings, and for geotechnical aspects of open pits, pipelines, railroads, roads, and so forth. We've also designed water supply dams, which could be considered recursors of tailings dams and other forms of disposal systems. Let me show by example what geotechnical engineering can contribute to waste disposal facility design. I recently was asked to look at several small wastewater impoundments which had failed. The design of the earth embankments appeared, at first glance, to be rather conservative. They were only about 3.4 m high, the slopes were 3 to 1 (horizontal to vertical), with about 24 m of bottom contact with the earth foundation. But there had been three episodes of dike failure in a few months, all upon initial filling. The problem was the nature of some material in the foundation that was also used as structural fill-a dispersive clay that readily deflocculated and eroded on contact with water, especially with seeping or flowing water. The design did not include the defensive measures required to use such materials confidently in water retention structures. Apparently, the designer did not recognize the potential problem posed by the unusual soil characteristics. That's the sort of engineering information fundamental to geotechnical engineering, but not necessarily a part of the civil, structural, or mining engineer's professional training or experience. This difference in expertise is not simply a matter of factual knowledge-it also affects the design process. The structural engineer can design a structure to fit the client's functional needs and can specify the materials (grade of steel, PVC, concrete, etc.) to meet the demands of the design. The geotechnical engineer's design process is somewhat different, since we work with natural materials, not man-made materials produced according to specification. We first have to establish what kinds of materials are available (usually directly from the chosen site, since importation of materials becomes expensive). Then we determine the engineering properties and characteristics and, finally, design the structure to fulfill the client's requirements within parameters dictated by available materials. In brief, the geotechnical engineer works with naturally-evolved, not man-made, materials and uses design techniques developed to accommodate that fact. Because no two sites are exactly alike in configuration, structural (geologic) characteristics, or available materials, there will be unique design and judgmental questions to be resolved on virtually every project. You said that water retention dams could be considered the precursors of tailings dams-you probably had a specific technical relationship in mind. There's a very strong carryover of methodology from earth dam design, where the technology developed and where it has become very refined. In designing a tailings impoundment or similar facility, essentially the same procedures of field sampling are involved (with a geotechnical engineer or an engineering geologist directing the program since representative

Jan 7, 1980

By D. A. Wheeler

Invented by Pierre Boutin in the early 1960s the column was a complete break from the conventional flotation cell. 1. When used as a rougher-scavenger, the column is excellent. 2. When used as a cleaner, the results can be spectacular. The very first column tests were carried out using it as a rougher-scavenger in a reverse float where silica was floated from iron. It produced a concentrate underflow as good or better than that produced from roughing and scavenging stages in cells, The froth tailing overflow was better than that produced by several stages of cell cleaning. Column scaleup progressed rapidly from the two inch diameter unit to a semi automated 12 inch diameter column on material from the Iron Ore Company of Canada. A change in operating philosophy within IOCC brought all flotation development, column or conventional, to a halt. At that time, IOCC had an exclusive right to the use of the column in Canada in the field of iron ore. We moved into the field of sulphides. A Canadian copper producer sent ore for testing and the results led to their purchase of the first commercial size of column - a 36 inch diameter machine. It was a mechanical disaster. It took several years to raise sufficient funds to return to that mill with our basic 18 inch square unit. It was to be tested and modified in order to learn how to properly design a large column. Originally used as a rougher- scavenger, it had to produce tails equal to the final tailings from this well run plant. It did, and did so while producing a rougher-scavenger concentrate almost equal to the plant final concentrate. It was finally used as a cleaner and produced concentrate 5% higher in copper than the plant final concentrate with equal cleaner tails. This phase of the column development was carried out under very difficult circumstances. The mill superintendent had realized that if he took the froth removal system from our original mechanical disaster and applied it to a conditioner while injecting air, he would have a flotation cell. He had personally applied for patents on this Maxwell Cell. Our development work was done in his mill and as our results became better and better, our difficulties became worse and worse. We finally had to terminate this work. The 18 inch column installed at Mines Gaspé 14 years later was identical to the one removed from Opemiska. The 18 inch column was tried on various ores over the following years and always produced excellent results. However, it was not really a production size unit. We had always aimed for the 72 inch column (72" x 72" x 44' 9"). Prior to this huge machine, we needed the intermediate 36 inch column. The failure of the original 36 inch diameter unit at Opemiska had raised the possibility of short circuiting inside the column as the cross section increased. The first 36 inch square unit was tested in parallel with the proven 18 inch column. If the underflow of the 36 incher was not as good as that of the 18 incher, short circuiting was a possibility in the larger unit. It had been designed for insertion of drop in partitions, four feed points had been provided and the 36 inch column would have become a modular unit of four 18 inch columns. Testing showed the underflows of both columns to be identical. There was no short circuiting in the 36 inch column. Once we had the 36 incher, we had no fear of the 72 inch column. It is permanently partitioned into four 36 inch columns but uses only one set of instrumentation. All the component parts of the 72 inch column come from the 36 inch unit. In spite of our results, the mining community did not believe the column could work. Finally, in 1980, Mines Gasp6 ordered an 18 and 36 inch column for their byproduct molybdenum

Jan 1, 1988

By V. A. Evans, D. L. Hellier

Introduction Preparing a report on fluorspar economics in the US could have been somewhat complicated at one time. This was because the number of producers, both large and small, derived production from different locations through-out the western US and the southern Illinois and western Kentucky fluorspar districts. Companies, such as Allied Chemical and Williamson Mining, operated in the Jamestown district of Colorado. In Walden and Salida, CO, the Colorado Fluorspar Co. operated mines and mills. In New Mexico, both Du Pont and Allied had mines and mills. In the Illinois-Kentucky district, US Steel, Rosiclare Lead and Fluorspar, Minerva Oil Co., and Marathon Oil all produced spar for the US market. Now all of the companies are out of the fluorspar business and, with the closing of the Seaforth Mineral Co.'s Inverness Mining Co., only Ozark-Mahoning Co. survives as a domestic producer. History Ozark-Mahoning Co., as it now exists, was formed in 1946 by the merger of Ozark Chemical Co. and Mahoning Mining Co. Ozark Chemical Co. was started in the 1920s to produce sulfuric acid for two refineries located in Tulsa, OK. Mahoning Mining was originally formed by the founders of Ozark Chemical Co. and the Youngstown Sheet and Tube Co. to produce fluorspar and associated minerals from properties in the southern Illinois and western Kentucky fluorspar district. In 1937, Mahoning Mining started an exploration program, employing contractor-operated churn drills to look for reserves in Hardin and Pope Counties in southern Illinois. Based on known reserves in the area and the results of the exploration program, work was started in the fall of 1938 to build a selective flotation plant at Rosiclare, IL. The mill was completed and put into operation in August of 1939. The company's first mining efforts were in the Cave-In-Rock district just north of Cave-In-Rock, IL. The deposits in this area were of the Manto or bedded type. They have been the mainstay of the company's operations in the Illinois-Kentucky fluorspar district. In 1950-51, the company expanded its operations by purchasing fluorspar properties from Colorado Fluorspar Co. at Walden, CO. This location was operated from 1952 through 1958 and from 1968 through 1973 when it was closed permanently. In 1974, Ozark-Mahoning Co. was purchased by Pennwalt Corp. Presently, it is a wholly-owned subsidiary of that corporation. As the only major domestic producer of fluorspar in the US today, the company's ability to survive under the flood of imported material is directly related to the four major segments of its operations. These are: exploration, mining productivity, maintenance, and marketing. All are done in-house. Exploration The Southern Illinois fluorspar district is located in the most complexly faulted area of the central US. The formations are sedimentary in nature and range in age from Devonian through Pennsylvanian. The deposits are small, they range in size from a few thousand tons up to 900 kt (1 million st), with the average minable deposit being around 544 kt (600,000 st). The ore occurs in both nearly vertical vein deposits and in Manto or flat lying bedded deposits. Due to the relatively small size of the deposits and the depth at which they occur, a good low-cost exploration program is essential to ensure that new finds of ore are equal to the mining extraction rate. Ozark maintains a leased land position in Pope and Hardin Counties that will fluctuate somewhere between a low of 40 km2 (10,000 acres) up to a high of 129 km2 (32,000 acres), depending on total size of tracts leased. Most all lands are obtained through a standard lease purchase option, with a small minimum royalty per acre during the normal 10-year option period. All leases are standard and without deviation to control land cost. All land leasing is done through Ozark's geological staff, who also

Jan 11, 1986

By N. L. Wesis

General History The first jaw crusher in the US was patented about 1830, but the Blake crusher that has maintained a substantial advantage over other types was invented in 1858 by Eli Whitney Blake. For going on 150 years the jaw crusher has been an invaluable machine; even today, when gyratory crushers have assumed most of the burden of crushing large tonnages of large rock and ore, the jaw crusher has a firm place in the mining industry. It is interesting that soon after the invention of the Philetius Gates gyratory crusher in 1881, a contest between the first No. 2 Gates and Blake jaw crushers of equal gape showed the gyratory to be 3.2 times as fast. This foretold a day not far in the future when the jaw crusher would not be adequate on the largest jobs. In his Textbook of Ore Dressing (1909) R. H. Richards recom¬mended that the word breaker be used for all machines breaking to relatively large sizes (say to 5 in. or greater), and crusher be used for finer work. Even in the 1968 US Bureau of Mines dictionary of mining terms the work breaker is defined by Richards in this context. However, in modern US usage it is nearly limited to coal breaking. Over the years the jaw crusher has been developed in a variety of forms but it appears today in these three general forms: (I) Blake type, (2) Dodge type, and (3) single toggle (or overhead eccentric) type. All of these crushers have a fixed jaw and a moving jaw, between which the coarse rock fragments are intermittently caught and crushed. The three types are differentiated by the manner in which the movable jaw is moved in relation to the fixed. Types The Blake-type crusher is shown in cross section in Fig. 1. The simplicity and strength that have made it first among jaw crushers are evident. The movable jaw is suspended from a cross-shaft at its upper end, this shaft being carried in bearings on the sides of the frame. The actuating mechanism consists of an eccentric shaft, also supported in frame-mounted bearings, which imparts through a pit¬ man and a pair of toggles a reciprocating motion to the bottom of the swing jaw, the return movement being effected by several spring¬loaded rods. The whole is built into a box frame with the crushing chamber at one end. As in all jaw crushers, the eccentric shaft is equipped with a heavy flywheel that maintains an even speed through¬out each stroke. The Dodge-type crusher illustrated in Fig. 2 is simple, mechani¬cally. Its movable jaw, being pivoted below the discharge opening, has minimum movement at crusher discharge and maximum at crusher feed. Because the choke point coincides with the point of least motion, these crushers are of relatively low capacity, and the rapid action in the feed area gives the machine the advantages of large reduction ratio and closely sized product. Thus, it fits well into low-capacity operations like sample reduction. Besides low capacity, its principal disadvantages are tendency to pack and make fines. The single-toggle or overhead-eccentric type of jaw crusher (see Fig. 3) has gained in usage, starting as a modification of the Dodge idea in that the greatest movement of the movable jaw occurs at the top, and being gradually improved and strengthened to its position today where it covers just as wide a field of application as the Blake type from the standpoint of feed opening. The motion of the movable jaw is the result of the circular motion of the eccentric shaft at the top of the swing jaw combined with the rocking action imparted to the bottom by the inclined toggle plate. Because of light weight and mechanical simplicity, it is economical in small mills and portable plants where the rock is not as tough and abrasive as to punish the machine with excessive impact from shocks, or excessive wear result¬ing from the pronounced vertical component of motion.

Jan 1, 1985

By Richard H. Kennedy, G. A. Swanquist, John W. King, Frank F. McGinley, Dale C. Mathews, Hans H. Adler, E. C. Peterson, Ralph M. Wilde

INTRODUCTION From the first historical records of the discovery of uranium as an element and the attempted classification of its minerals in the early 18th century-through its first uses in pigmentation and the coloring of glasses and ceramic glazes-until 1939 when it was discovered that the isotope U235, was fissionable, uranium oxide in the form of pitchblende ores and concentrates (produced mainly in Bohemia, Canada, and the former Belgian Congo) in recent times was the source of uranium compounds incidental as byproducts to the recovery of radium from this invariable associated source. In 1938 this byproduct production from pitchblende concentrates was about 471 st of U3O8, with over 90% from Canada and the Belgian Congo and only 25.85 st from carnotite ores in the United States.1 Under the impetus of the Atomic Energy Commission's (AEC now part of the Energy Research Development Administration) domestic and foreign production incentives and buying programs, and as an initial assurance of atomic armament capability of the United States and the Free World, the growth of the uranium mining and milling industry burgeoned from 1952 to 1962. After a short period of decline, in the second decade and in large commitments beyond, it has again extended its capacity to meet ever-increasing fuel demands of atomic power generation by the electric utilities industry. In 1961 the peak production in the United States alone was 17,758 st of U3O8. Peak aggregate domestic and foreign deliveries of 34,581 st of U3O8 to the AEC occurred in 1960.z Canada's maximum delivery of 13,506 st of U308 was made in 1959. These peak and continued productions in the Free World are largely from: the Mesozoic (Jurassic) and Tertiary (Eocene) sandstones in the United States; the quartz pebble conglomerates of Canada and the Witwatersrand; pegmatites, Jurassic limestones, and Precambrian complexes, with less but significant production; the phosphoria, where byproduct production is minor. This uranium production in the form of precipitated concentrates has come from the milling of ores from these sources, both directly and by retreatment of gold cyanidation tailings on the Witwatersrand. Milling of crude ore has followed these processes: conventional crushing and size preparation, ore drying when required, salt roasting for high vanadium ores and byproduct vanadium recovery, carbonate or acid-leaching processes, classification sand-dine separations or countercurrent decantation (CCD) circuits and complete liquid-solids separations, resin-in-pulp (RIP) or liquid-liquid (solvent) ion-ex- change extractions, clarification of pregnant liquors, precipitation, filtration and washing, sodium and vanadium removals, and drying or roasting and packaging of the high-grade, high-purity uranium concentrate commonly referred to as "yellowcake." Coupled with the hazards and control of dust generation incident to conventional crushing and milling operations, more stringent control of airborne respirable and ingestible radioactive particulates and the protection of personnel against definitive external radiation are inherent to uranium milling. Impoundment and control of tailings and other effluents of uranium milling are more restrictive than in conventional practice, be- cause long-term disposal and containment of the uranium series of radionuclides, whether in solution or in solid form, are requisite to uranium milling and are made manadatory under regulatory agencies of government. 2. GEOLOGY AND MINERALOGY OF URANIUM Early Uses and Byproduct Production of Uranium Uranium was discovered by Klaproth in 1789 but had little commercial importance until the discovery of uranium fission by Hahn and Strassman at the close of 1938. The well-known uranium ores of Joachimsthal, Czechoslovakia, the Congo, and the Colorado Plateau were early sources of radium. Minor quantities of uranium were used for small-scale technical and industrial applications, particularly in coloring glass and ceramics, but most of the uranium was discarded. The Colorado ores were also mined later, chiefly for vanadium. Impetus of Armament Capability and Fuel Demands for Atomic Power Following the discovery of atomic fission, uranium became critically important for military applications and, more recently, as a nuclear fuel. Prior to 1966 uranium was acquired in the United States almost solely for use in the weapons program of the AEC. By 1970 the AEC had procured approximately 325,000 st of U3O8 from domestic and foreign sources. Current and future marketability is primarily dependent upon commercial demand for reactor fuel. Estimates of

Jan 1, 1985

By Theodore M. Gurr, James J. Bartels

Phosphate produced fertilizers provide the phosphorous nutrients required by plant life to sustain significant growth, thus improving the production of food for the world's population. Phosphorous replenishment of soil currently can be obtained efficiently by no other means than by direct application of phosphate fertilizer to the soil. Phosphate products are also utilized in animal feeds, detergent, and various industrial processes. Thirty-three countries are presently mining phosphate rock for the production of fertilizers, and its process byproducts. The major phosphate rock producers are the United States, the former Soviet Union, Morocco, and China [(Table 1)]. Phosphate rock is generally defined as a rock material that contains phosphate minerals which are sufficient for commercial usage. Phosphate minerals are found in sedimentary igneous and metamorphic rocks. Economic resources of phosphates are primarily developed from sedimentary rock sources. The apatite mineral family comprises the majority of phosphate constituents from both sedimentary and igneous rocks. Igneous phosphate rocks are generally composed of fluroapatite whereas sedimentary rocks are generally composed of carbonate fluroapatite, or chlorapatite. Phosphate rock is mined by a multitude of processes, including surface and underground mining. The mined rock requires mechanical and chemical processing to liberate the phosphate for utilization as fertilizer. Approximately 85% of the world's phosphate production uses sulfuric acid, with the remaining processes utilizing nitric and phosphoric acids. Ammonia is also introduced in the manufacture of liquid fertilizer, ammonium phosphate and ammonium polyphosphate. GEOLOGY Mineralogy and Chemical Properties Phosphate most commonly is derived from the mineral apatite, which is chemically described as CA5(F,Cl,OH)(PO4)3. Apatite is also broken up into the composite apatite minerals which are fluroapatite Ca5F(PO4)3, chlorapatite Ca5Cl (PO4)3 and hydroxyl-apatite Ca5(OH)(PO4)3. Carbonate CO3 can substitute for PO, forming carbonate apatite, called francolite. The guano phosphate (bird excrement) mineralogy commonly occurs as brushite CaHP04.2H20, monetite CaHPO4, whitlockite Ca3(PO4), and dahlite Ca10(PO4)6(OH)2. Apatite minerals form from igneous sources where they are developed deep in the earth's crust; and the cooling processes are slow, producing long, hexagonal prismatic to tabular crystals. The termination of the crystal can be basal plans or pyramids. In some instances, the crystals are bipyramids. Faster cooling, or sedimentary derived apatites are more commonly cryptocrystalline. Collophane is a name given to cryptocrystalline apatite found in phosphate rock in fossil form. Apatite minerals are relatively soft and are used to describe the 5 hardness on the Mohs scale. Cleavage is poor and develops along the C(0001). The color of apatite minerals varies from colorless, to violet, and blue, but are predominantly green or brown; luster is vitreous to subresinous. Collophane's physical appearance is often opaline, dense, with colloform structure and sometimes concretionary, nodular or pulverulent. Phosphate Bearing Materials Origins Phosphate deposits of igneous and metamorphic origins have been well defined. Sedimentary origins have been defined, but with a greater degree of uncertainty. Research in the 1980s proved out some of these processes through studies of real time deposition in Chile, Peru, and Australia. Igneous intrusive alkali rock and associated contact metamorphic rocks, provide approximately 20% of the world's phosphate. Fluorine containing apatite minerals are the most common materials containing phosphorous. Fluorapatite occurs in most igneous rocks. Fluorapatite and fluorine hydroxylapatite, together with carbonate varieties of these, are important members of the group. Where an essentially pure chlorapatite, carbonate apatite, and hydroxylapatite are rare and restricted in occurrence, the fluorine containing types occur in most all igneous rocks as early formed accessory mineral, usually in microscopic crystals, and may occur as extremely large bodies as magmatic segregations from alkalic igneous rocks. Apatite is also found crystallized in pegmatitic faces of both acidic and basic types of igneous rocks. More specifically, apatite is associated with magnetite deposits, in hydrothermal veins, especially those formed at relatively high temperatures, and in veins of the Alpine type. Apatite is common in both regionally and contact metamorphosed rocks, especially in the crystalline limestones where it is associated with sphene, zircon, pyroxine amphibole, spinel vesuvianite, phlogopite, talc, chloride schists, and as a contact metamorphic mineral.

Jan 1, 1994

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By Robert W. Miller, Rhoda S. Kriesel

INTRODUCTION With each passing year, the general public becomes increasingly aware of the potential hazards associated with many products and services previously considered unharmful. This new consciousness is especially evident in the workplace, both as the result of advances in scientific understanding and the influx of concerned, responsible professionals into the labor movement. For example, data produced by epidemiologists doing retrospective mortality studies which related mortality to toxic substance exposures has often been the prime motivating factor behind reductions in exposure standards. This certainly has been the historical case with many radiation exposure standards, including the limitations set for "Working Level" exposure. The concept of "Working Level" was introduced in PHS Publication No. 494 as a result of both the difficulty in relating radiation units to biological effects and the complexity of interaction of radon daughters with the physical environment. WORKING LEVEL One working level is defined as any combination of radon daughters in one liter of air that will alternately release 1.7 x 105 MEV of alpha energy during decay of 210pb (RaD). Its usefulness is that working level can be readily measured in both the field and the laboratory. Because of this, existing occupational exposure standards for radon daughters in both the U. S. and other countries use working level as their basis. In the U. S., the last twenty years has witnessed a significant reduction in radiation exposure standards from 120 working level months to 4 working level months per year. New international standards are being proposed that would be only slightly higher than these current U. S. standards. POTENTIAL EXPOSURES TO RADON DAUGHTERS The universal use of nuclear power, along with the transfer of nuclear technology to developing nations, has significantly increased the demand for fuels. As a result, uranium mining operations have expanded and increased numbers of miners are regularly exposed to radon daughters. In addition to potential exposures in uranium mines, other deep mines such as tin, gold, platinum and tungsten, also have potential exposures due to the distribution of uranium and its accompanying radium throughout much of the earth's crust. The energy crisis has also prompted interest in general population exposure to radon daughters in buildings. Many homeowners, using additional insulation and caulking to seal cracks and conserve energy, have reduced indoor ventilation to the point where working level exposure could become significant, especially in homes with unventilated crawl spaces. In fact, several serious exposure situations have occurred where mine tailings were used for building materials. MEASUREMENT TECHNIQUES Historically, rapid and convenient measurement techniques have been actively sought to improve the ability of ventilation engineers to limit mining exposures through ventilation control. A number of measurement techniques have been developed, including the Kusnetz method (further elaborated by Shalaynev) and techniques developed by Rolle and Tsivoglou. All, however, suffer from similar handicaps that prevent their usefulness in the mining environment. Their main limitation, in terms of minimizing employee exposure, is the elapsed time from the start of sampling until the results can be calculated. It is usually 40 to 90 minutes. This time delay presents both serious economic limitations as well as the potential for exposing workers to high radon daughter concentrations until calculations are completed and corrections made. Further reducing the utility of the Kusnetz and Tsivoglou techniques is the need for tedious calculations that increase the possibility of human error along with the need for cumbersome sampling equipment that is not ideally suited to the mining environment. INSTANT WORKING LEVEL METER To alleviate these problems, the idea of an Instant Working Level Meter was conceived by several groups. Several such instruments were proposed, built, and tested with disappointing results. Although the concepts behind them were sound, the instruments simply failed to measure working level without excessive distortion. In addition, their cost was relatively high, gamma background presented a problem, and sophisticated computations were necessary to determine working level. In short, these systems lacked the design engineering that would make them suitable for the demands of a mining environment.

Jan 1, 1981

By Edward Burris, Terry Gorsuch, Joseph M. Hans

INTRODUCTION In any large earth-moving operation, good planning and management are necessary to complete the operational tasks promptly and successfully. When an earth moving operation is complicated by radioactive contaminants, normal earth moving techniques and procedures must be modified. Any planning and management, therefore, must include the radiological aspects of the operation. It was found that the radiological aspects dominated most of the planning and management activities and were extended to all facets of the decontamination work at the former Shiprock uranium millsite. These planning aspects are discussed and their use to develop a work plan is described. The management aspects are discussed and their use to establish a management structure are also presented. PLANNING Some method of procedure, formulated beforehand, was necessary to govern the decontamination work at the former Shiprock uranium millsite. This procedure was expressed in the form of a work plan which served several listed purposes. 1. It defined the work to be done and the sequence it would follow. 2. It was used as a yardstick to measure progress. 3. It was used to assign organizational responsibilities. Several factors were considered to aid in the development of the plan. These factors are discussed below: Goals It was established that radiation exposure was occurring to persons working at the millsite, and in an around the community of Shiprock, from airborne radioactive mill wastes and radon-222 exhaling from the tailings piles. The goal set for the decontamination work was to reduce on-site exposures to levels acceptable for the millsite occupants. The attainment of this goal would also have a substantial impact in reducing off-site exposures. The objectives necessary to achieve the goal were consolidation and containment of the wastes. The former objective implies decontamination of the millsite and environs, and the later implies stabilization of the wastes. In practice, a total and complete decontamination of the millsite and contaminated environs would be very difficult and costly. The costs for decontaminating them could be high enough that an alternative method might be more cost-effective for reducing human exposures (i.e, move the affected people away from the source). The interim guide "Radiological Criteria for the Decontamination of Inactive Uranium Millsites" was used for the decontamination criteria (EPA 74). Briefly, the criteria state that off-pile decontamination should be effective enough to reduce the net above ground exposure rate to less than 10 [u]R/hr for unrestricted use of the affected area. When decontamination cannot readily be achieved, the exposure rate levels could be relaxed to 40 [u]R/hr; however, the affected area has to be restricted. The second objective, waste containment, means isolating the wastes from the biosphere. Since no method of containment was available at the beginning of the millsite decontamination effort, temporary containment (interim stabilzation) became the objective. The tailings pile and decontamination wastes would be covered with clean fill. The interim stabilization should last from 5 to 10 years until the final disposition of the wastes will occur. The goal, therefore, would be achieved by decontaminating the off-pile areas to less than 10 uR/hr where practical. The decontamination wastes would be used to plate the surface of the tailings pile and would be covered with clean fill. Radiological Survey The radiologial survey is the key factor for planning a decontamination activity. The survey should delineate the spread and depth of the contaminants relative to the decontamination criteria. Surface wastes, in general, can be evaluated for spread and depth with reasonable radiation survey equipment. Subsurface wastes on the other hand can be missed entirely, as happened during the radiation survey at the Shiprock site, although numerous exploration holes were bored and dug. The survey results can be used to define areas that may not be amenable to decontamination because of complications or safety reasons. For example, no decontamination of the bluff base was to be attempted because of the possibility the bluff might collapse on the personnel and equipment. Contaminated bottoms of decant ponds on the flood plain were not removed because they would be slurried by ground water. Slurry removal was deemed inefficient because the contaminants would be scattered and no equipment was available for its transport. In summary, the radiological survey defines the boundaries of the decontamination work and provides

Jan 1, 1981