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By Gaither H. Kitchell

Introduction Fire assaying is an ancient method of determining the amount of gold and silver in geological materials. Today's commercial, mining, school, and private fire assaying laboratories still use this basic assaying method. One of the primary reagents used to collect gold and silver from the ore is litharge (PbO), or lead oxide. The Federal Occupational Safety and Health Administration's (OSHA) and the Nevada Division of Occupational Safety and Health's (OSH) have a lead standard. It states that a person exposed to airborne lead at concentrations greater than 50 µg/m3 (1.4 µg per cu ft) of air over an eight-hour period has gone beyond the permissible limit for health safety. It is not uncommon to get airborne lead concentrations of 1400 µg/m3 (40 µg per cu ft) in the flux/sampling area or 900 µg/m3 (25 µg per cu ft) around the gas/ electric furnaces during the fusion and cupelling phases. Therefore, any individual working in these areas for more than 30 days per year without proper protection is a likely candidate for lead poisoning. Lead In A Human Lead enters the body as a toxic substance by breathing it into the lungs and upper respiratory tract and by getting lead particles into the digestive system through ingestion. Smoking, eating, drinking, chewing tobacco and gum, or wiping a lead-contaminated hand across the mouth are all potential ways forgetting lead into the body. Lead taken in large quantities can actually kill a person in a matter of days. This is rare, however. The real hazard of lead overexposure is for the unsuspecting person who works in an unsafe lead environment for months or years. This individual can slowly receive damage, sometimes permanent, to the central nervous system, gastrointestinal tract, kidneys, blood, gums, and reproductive system. Safe Blood-Lead Level The Lead Standard states that the maximum safe blood-lead level is 40 µg/100 g (40 µg per 3.5 oz) of whole blood. It recommends 30 µg/100 g (30 µg per 3.5 oz) if a person is of child-bearing age. People not working around lead products normally have a blood level of less than 10 µg/100 g (10 µg per 3.5 oz). A person not adequately protected can exceed 40 µg/100 g (40 µg per 3.5 oz) in a few weeks by fire assaying only 100 samples per day per week. Studies show that some individuals have developed lead-related diseases with blood-lead levels less than 80 µg/m3 (2.3 µg per cu ft). Lead Poisoning Symptoms There is a long list of symptoms associated with lead poisoning. The most common are: a prolonged, sweet, metallic taste in the mouth; loss of appetite; anxiety and irritability; weakness and dizziness; nausea and tremors; and insomnia. Preventive Measures Based on the lead standard, a number of preventive measures can be taken to ensure that fire assay laboratory personnel frequently exposed to lead are protected against overexposure. First, the actual lead concentration within the assaying area should be determined. In Nevada, one can contact the Department of Industrial Relations, Division of Occupational Safety and Health for assistance in monitoring any laboratory. The agency will also recommend ways to protect against lead. This agency does the monitoring and provides an educational class on lead safety. This type of agency can be found in other states. Commercial consulting firms are also available, but they are expensive. Regardless of the technique used to determine lead concentrations, it is most likely that the average fire assay area exceeds the 50 µg/m3 (1.4 µg per cu ft) lead concentration limit. This is considered unsafe.

Jan 8, 1985

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

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

Jan 1, 1986

By J. B. Gwiazda

In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the .u factor used by designers of lemniscate-guided roof supports has never really been confirmed as a maximum and assumes that convergence is vertical. Also, the paper does not appear to take into account deflection of structures, which occurs when the lemniscate and base members are fully loaded to their maximum stress level, nor the front to back line of the support in relation to differential roof to floor movements caused by strata movements under pressure. It is not unusual for differential movements to be slightly diagonal to the line of the support, particularly in faulted areas and on gradient faces. The paper also does not take into account consolidation of fines immediately above and below the support. Generally speaking, any differential movement is from face to waste and under these conditions the .u of 0.3, which appears to be an international standard, has worked in practice. However, if the face end of the support is lower than the waste end, then the µ of 0.3 can be considerably increased, giving rise to the damage mentioned in the paper. The ideal design should aim for a slightly forward bias in the lemniscate guide so that the last increment of setting is toward the face, tending to close any fissures that may have developed during the support advance cycle. The support should also be fitted with positive set valves to ensure that a high setting load density is attained to minimize bed separation. As far as powered supports are concerned, convergence is irresistible and all powered supports converge at their rated yield load. A similar principle can be applied to the differential roof to floor movements to drastically reduce the very high forces that would otherwise be applied to the lemniscate structures and pins and that, in turn, are transferred to the base arrangement and floor loading. Any differential movements are usually catered for by the 0.3 µ factor or deflection of structures in the lemniscate guide arrangement and consolidation of the floor. The floor loading, due to differential movement, is in addition to the support convergence load and requires additional bearing area to avoid possible floor penetration. Some seven years ago, Fletcher Sutcliffe Wild Ltd. (FSW) introduced a lemniscate-guided shield support where the lemniscate linkage is connected to the roof bar through two horizontally converging rams to allow differential movement to take place above a given rated figure. This is a known force and can be guarded against, whereas with rigid connections the forces, as yet, are unconfirmed. By careful design, a horizontal force in excess of 6 MN (60 tons) opposes differential movements for a total ram loading of only 2.5 MN (25 tons), or 1.25 MN (12% tons) each. This principle can considerably reduce the length and weight of the support in comparison with a rigid pin-type structure; also, the yield load rating can be increased without affecting the lemniscate forces. The graph shows the tensile and compressive forces in a lemniscate linkage of a support with and without hydrostore. These forces react into both the roof beam and base members and, as can be seen from the support height to linkage load graph, a considerable reduction in these reactions is gained by the use of the FSW patented hydrostore system. Floor loading is considerably reduced under maximum µ conditions, and by allowing the roof bar to move with the strata, some degree of improvement to strata control is achieved in line with the assumptions in the paper. In practice, these movements have only been in the region of a few millimeters, which, in turn, reflects on the improvements to strata control by the addition of positive set valves.

Jan 1, 1987

By M. Ge, C. Mborah

"The acoustic emission/microseismic monitoring (AE/MS) technique has widely been used in the mining environment for the monitoring of rock mass stability. A successful implementation of the method is hinged on being able to accurately determine the P- or S-wave arrival which is then used in the determination of the source of an event. Determination of the arrival phase is usually performed visually by a human expert. However, the mining environment is characterized by high noise levels and different noise types of varying intensities. Therefore, performing manual phase picking on time series data acquired from the AE/MS monitoring system can sometimes be tedious and time consuming if not impossible. In this paper, a proposal for improving manual P-wave phase arrival picks is discussed. The method involves the filtering of the AE/MS data using a stationary discrete wavelet transform (SDWT) method. After filtering, appropriate scales/levels of the multi-level decomposition process are then selected for further processing. The first step in the picking process after filtering, requires the computation of the signal power and the root mean square (RMS) using the signal amplitudes at each scale. The preliminary P-phase arrival time is picked using the results of the calculated power and RMS values respectively. The calculated RMS values are then superimposed on the calculated power values to determine the final arrival picks. If the arrival time picks obtained by both methods coincide or are within 5 to 10 sample points, the pick is considered reliable. The accuracy of the method was verified using AE/MS data from two datasets obtained from two separate underground mines. The results of the study showed that the picking method is reliable and robust. Based on the data studied, it can be concluded that the method will help reduce the time usually spent in processing AE/MS data from mines. INTRODUCTION A successful identification and location of the source of acoustic emission/microseismic (AE/MS) events is mainly dependent on the onset time of the P-wave phase. In many seismic networks, initial picks are performed by picking algorithms followed by manual picks by human experts. If the waveform signal-to-noise ratio (SNR) is low, the P-wave onset time picks by the human expert is generally reliable and accurate [1]. However, for waveforms with high SNR, picking of the P-wave onset could be a major challenge. Also, due to the volume of data acquired daily, manual processing could be stressful and time consuming [2]. These challenges continue to provide motivation for alternative means of providing reliable and efficient ways of improving manual phase picking. The AE/MS technique has been an intergral part of safety monitoring in mines [3] due to its capacity to identify zones of instability in a monitoring area. To obtain an accurate source of an AE/MS event however depends on the quality of the data used for P-wave phase picking [4, 5]. AE/MS data from mines are generally contaminated with excessive background noise generated by the different mining operations. The AE/MS signals of interest are mostly buried in background noise which makes it difficult to pick the P-wave phase arrivals [4]. Processing AE/MS data acquired in noisy environment remains a major challenge due to the weak nature of AE/MS signals [6]. As noted by [4] most of the data recorded in the mine environment are mainly due to events not related to AE/MS events of interest. The complexity of data in this type of environment makes the detection and picking of P-wave arrival by human experts a very difficult task."

Jan 1, 2017

By Ernest E. Thurlow

Introduction Today, more than 50% of rail-carried commodities are mineral industry related, with coal being the most important single commodity moved by rail. In 1980, coal accounted for more than 5.7 million of the total 22.6 million carloads moved by rail. Metallic ores were third behind grain, with more than 1.4 million carloads. Crushed stone, gravel, sand, and other non-metallic minerals totaled almost 2.3 million carloads. Chemicals and allied products, including fertilizer and coke, added another 1.7 million cars, with petroleum and petroleum products totaling 300,000 carloads. Coal Several things happened in the 1970s that gave rise to increased consumption of coal-particularly western coal-and to its dominant position among rail-carried commodities. First, the Clean Air Act of 1970 required generating plants to make significant reductions in sulfur dioxide emissions. To comply, utilities could either invest in scrubbers or switch to low-sulfur western coal. Many opted for the latter. By 1972, increasing demands by the utility companies halted what had been a 25-year decline in national coal consumption. Second, the Arab Oil Embargo of 1973 put an end to cheap oil and gas, limiting their future as fuels for electric generating purposes and increasing the potential for coal. Third, in 1977, President Carter announced an energy program with coal as its cornerstone, calling for an annual two-thirds increase in national coal production by 1985. He also called for conversion to coal by utilities and large industrial users. Finally, he proposed a 10-year, $10 billion program to encourage domestic coal production and stimulate development of export markets. Coal bounded into world prominence. Foreign demand for steam and metallurgical coal increased tremendously, while US demand for western coal also shot up. This meant greater demands on the transportation sectors that traditionally carried coal to market. Many railroads began programs to serve the coal industry. One example is Burlington Northern's commitment to handle increased western coal tonnages. The company spent more than $1 billion in recent years to develop a system capable of moving more than 91 Mt (100 million st) of coal each year. Other leaders in this renewal were Norfolk and Western, Union Pacific, Santa Fe, and Southern Pacific railways. The importance of the rapid growth of coal traffic to the railroads is shown in the accompanying table, which gives percentages of total tonnages hauled and revenues attributed to coal. With coal providing the railroad industry with a substantial share of its revenues, there is keen competition among the rail companies themselves and among railroads and other transportation sectors for coal haulage. But there is also cooperation when more than one railroad is involved in delivering the coal from mine to market or when a combination of transportation modes is more economical. The latter is represented by Conrail's interest in working with the port authorities of New York and New Jersey to establish a new coal port that would serve not only export markets, but also utilities and industries in the northeast. Iron Ore Next to coal, iron ore (taconite) is the most important single mineral commodity handled by railroads. In Minnesota, where most iron ore is produced, rail transportation is primarily by Burlington Northern and the Duluth, Missabe, & Iron Range Railway Co. (DM&IR). The DM&IR, owned by US Steel, serves several producers on the Mesabi Iron Range. Two of the larger producers, Erie Mining Co. and Reserve Mining Co., also own railroads that operate between the mines and the ports of Silver Bay and Taconite Harbor on Lake Superior. Several iron mines and taconite plants in Michigan are served by the Chicago and Northwestern, and the Soo Line. Total 1981 shipments of taconite and iron ore are estimated at 55.9 Mt (55 million It), compared with about 61 Mt (60 million It) in 1979 the most ever shipped in one year. Still, with annual production capacity of the eight Mesabi Range

Jan 10, 1982

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 Werner Hofmann

INTRODUCTION The attempt to interprete lung cancer incidence caused by inhaled radon decay products in items of radiation dose raises the question of the relevant dose. The various dosimetric approaches published by now are characterized by an increasing complexity in the structure of the models used. Ranging from organ doses to microdosimetric specific energy distributions in the cells at risk, a significant trend from the macroscopic to the microscopic scale can be observed. Since tumors originate preferentially in the basal cells of the lobar and segmental bronchial epithelium, these cells are regarded as the critical target for inhalation carcinogenesis. The evaluation of basal cell doses is, therefore, the main objective of lung dose assessments. Various dosimetric models have been published in the past (Altshuler, Nelson and Kuschner, 1964; Haque and Collinson, 1967; Harley and Pasternack, 1972; Hofmann, Steinhäusler and Pohl, 1979, 1980; Jacobi, 1964; Jacobi and Eisfeld, 1980; James , Greenhalge and Birchall, 1980). Due to incertainties of physical and biological parameters involved and the application of different anatomical and physiological models the claculated doses are varying up to one order of mangitude. Independent of the choice of these dosimetric models applied to the calculation of lung doses, deposition, clearance and resulting dose are affected essentially by the properties of the inhaled aerosol as well as by personal characteristics. Both the properties of the exposed individual as well as of the inhaled environmental atmosphere have to be considered in radiation protection. It is the main objective of this paper to illustrate the influence of both parameter sets on basal cell doses by varying the above parameters. PERSONAL CHARACTERISTICS Though it might be more convenient for radiation protection calculations, the biological "man" cannot be represented by a uniform standard man. In Tab. 1 these personal characteristics which cause the variability of man are listed together with the region of their effectiveness. In the following the influence of these parameters will be discussed with regard to lung dose assessments. Besides these factors the great biological variability of all anatomical and physiological parameters involved as well as pathological alterations due to lung diseases cause additional variations in lung dosimetry. [ ] Age Age can be considered as the most important personal factor influencing the anatomical structure of the lung between birth and about 15 years of life. This period is characterized not only by an enlargement of already existing respiratory units but also by structural changes. Besides these anatomical alterations deposition is strongly dependent also on respiration parameters, such as tidal volume (TV) and respiratory frequency (RF). Whereas RF decreases with progressing age in an exponential manner, TV increases significantly resulting finally in a growth of the respiratory minute volume (RMV). Also mucus clearance in the infant might be more effective than in the adult, though no reliable experimental data are available. Superposing all these age-dependent changes mean basal cell doses are higher for children by about a factor of 2 than for adults living in the same radioactive environment (Hofmann, 1981). In the adult the following alterations of respiration standards with age are reported: total lung capacity diminishes with age, functional residual capacity is unchanged or increases with age and the dead space rises slightly with age (ICRP, 1975). The influence of these lung capacities on deposition, however, can be regarded as insignificant. Due to the ageing process mucus velocities my be considerably smaller in older people. This effect can also be superposed by the simultaneous influence of ciliatoxic effects of cigarette smoke and air pollutants or other bronchial diseases. Since the lower age limit for occupational exposure is about 18 years, age can be disregarded in the uranium mining industry for the determination of radiation doses, though lung cancer induction is significantly dependent on age.

Jan 1, 1981

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 H. Polat, Q. Hu, M. Polat, S. Chander

The electrostatic charge on spray droplets of ionic surfactant solutions and coal particles was measured and the results were correlated with the dust collection efficiency. When various surfactant were added, the magnitude of the droplet charge increased significantly and it was observed to be a function of surfactant type and concentration. The concentration of maximum droplet charge coincided with surfactant concentration where maximum collection efficiency was observed for these surfactants. Particles of coal also carried substantial amount of charge magnitude of which seemed to be a function of coal rank. Based on the results presented in this paper, it was concluded that ionic surfactant primarily act as a strong electrostatic charge inducer for droplets. Due to interactions between these highly charged droplets and naturally charged particles, the efficiency of droplet-particle collisions play a primary role when compared to the wetting and engulfment phenomenon which could only follow a successful collision. INTRODUCTION Water spray are widely used to suppress airborne dust in mine atmospheres (Walton and Woolcock, 1960; Kobrick, 1970; Hamilton, 1974; Jayaraman et al., 1986). Several investigators have considered the use of surfactants to enhance the effectiveness of water sprays especially for difficult to wet particles such as those of coal (Glanville and Wightman, 1979). The capture of dust particles by water droplets involves droplet-particle collisions, adhesion of particles to droplets, and engulfment of particles into droplets. Surfactants affect these sub processes through their influence on droplet charge, surface tension, and wetting. The last two mechanism have been thoroughly studied in recent years (Walker et al., 1952; Cohen and Rosen, 1981; Glenville and Haley, 1982; Chander et al., 1988; 1991). However, little attention has been paid to the role of electrical charge on particles and droplets on the collision and adhesion of spray droplets and dust particles. Airborne particles of dust have long been known to carry a significant amount of electrostatic charge (Hopper and Laby, 1941; Kunkel, 1948; Kunkel, 1950; Dodd, 1952; Liu et al., 1987; Kutsuwada and Nakamura, 1989). It is reasonable to assume that presence of charge on particles will effect their agglomeration and particle-droplet interactions. Polat et al., (1991) showed that virtually all freshly generated dust particles were agglomerated in air. They suggested that electrostatic charge and humidity were important factors responsible for agglomeration. Previous theoretical studies on the interactions between charged particles and collectors by Nielsen and Hill (1976) show that, the collision efficiency is a strong function of the particle charge. In addition to charge on particles, spray droplets might also carry substantial amount of charge (Chapman, 1937;1938; Blanchard, 1958; Iribarne and Mason, 1967; Jonas and Mason, 1968; Byrne, 1977; Bailey, 1988). In theoretical studies of interactions between a spherical collector and airborne particle, it was found that the collision efficiency was significantly altered depending on whether the collector and the particles were charged. If neither the collector nor the particles carried a charge, the collision occurred by inertial and gravitational forces. The collisions took place on the front part of the collector (the front capture). If either of the collector or the particles were charged, the collision was enhanced due to the induced image forces. If both the collector and the particles were charged, the collision efficiency was significantly affected by the sign of the charge as well as its magnitude. For oppositely charged collector-particle pairs a collision could take place on the rear of the collector even if the particle flied past the collector upon approach (the rear capture) (Nielsen and Hill, 1976; Wang et al., 1986; Chang et al., 1987). On the other hand, the columbic force became negligible as the particle size increased and the inertial force became dominant. The electrostatic attraction was predominant for particles of less than about 2.5 µm in diameter. For particles larger than about 8 µm the inertia of particles was sufficient to overcome the columbic force and inertial impaction became the dominant collision mechanism (Grover and Beard, 1975; Chang, 1987). Previous studies of dust suppression using charged spray droplets generated by applying high voltage to the spray nozzle showed significant improvements in collection efficiencies (USBM open file report, 1983; McCoy et al., 1985). However, it was considered that highly charged spray droplets obtained by direct charging might have

Jan 1, 1993

By V. B. Achanti, A. W. Khair

Introduction Laboratory studies were carried out at West Virginia University to investigate the rock-fragmentation mechanism of continuous miners using an automated rotary cutting simulator. The primary factors influencing the fragmentation process were found to be bit spacing, bit geometry, depth of cut and cutting-drum rotational speed. This paper presents a discussion of the effects of these parameters in achieving optimum energy consumption and minimizing dust generation during rock fragmentation. The removal of rock ridges/walls between adjacent grooves is analyzed with three hits mounted simultaneously on the cutting head, while the bit tip angle was varied from 600 to 900. Bit spacing was varied from 25.4 to 50.8 mm (1 to 2 in.) while the cutting process was assessed for varying cutting depths. Respirable dusts generated during the course of the experiments were analyzed utilizing cascade impactors. Assessment of these parameters has led to a better understanding of the cutting mechanism of continuous miners in terms of energy consumption and dust generation. A review of the literature revealed that a considerable amount of research has been carried out on rockcutting processes. Many authors agree that the mechanical cutting efficiencies of mining machines (e.g., continuous miners, shearers and road headers) are affected by a host of parameters. Some of these parameters are machine controlled, some are operator controlled, while others are uncontrollable. Efforts were focused on understanding the influence of parameters such as bit spacing, cutting depth, attack angle, bit type, drum speed, bit geometry (i.e., tip size, shape and tip angle) and rock type on the cutting process efficiency in terms of specific energy consumption and respirable dust generation (Strehig?? et al., 1975, Hanson et al., 1979, Khair et al., 1989). Roepke et al. (1976) in an attempt to study the dust and energy generated during coal cutting using point attack bits found that the dust and the specific energy consumed both decrease as the depth of cut increases. The four fundamental stages of dust generation luring rock fragmentation are identified by Zipf and Bieniawski (1989). Coal breakage by various types of wedges was assessed by Evans and Pomeroy (1966) in an extensive experimental study on British coals. Yet the industry today requires further attention and guidance to optimize the energy consumption and dust generation during the rockbreakage process. This paper attempts to give a better understanding of the influence of some of the parameters listed above and focuses on further improvement in the rock-cutting process. The specific energy consumed for different types of bits used and the respirable dust generated are analyzed in the context of the variation of a few other parameters. Laboratory investigation The experiments were performed in the Rock Mechanics Laboratories located at West Virginia University. A rock-cutting simulator designed and fabricated by Khair (1984) was utilized for this purpose. The details of this machine are available in the literature (see Khair 1984). For this study, a series of preliminary experiments was carried out to determine the optimum cutting frame advance speed. This was intended to facilitate a maximum cut depth of 31.75 mm (1.25 in.) at an advance rate of 0.525 mm/s (0.0207 ips) for all types of bits being used and various bit spacings being considered. To look into the cutting-process efficiency of a continuous miner in the laboratory, several parameters of influence are being considered. Besides the bit-geometry parameters, machine- and operator-controlled parameters, such as spacing of bits on the cutting head, the cutting head rotational speed and the total cutting depth during an experiment, are varied. At the time this paper was written, only part of the completed experiments were ana¬lyzed, and a number of experiments were still being carried out following an orthogonal fractional factorial experimental plan to assess the effect of all of the above¬mentioned parameters on the cutting efficiency in terms of energy consumption and dust generation. Three different types of tip angles, namely, 60°, 75° and 90°, and two different tip sizes, namely, diameters of 7.94 and 24.61 mm (0.313 and 0.969 in.), were used. At

Jan 1, 1999

By A. Hayatdavoudi, E. Apande

Introduction Troublesome shales have plagued the petroleum industry for more than 50 years. Unstable boreholes have been experienced when drilled into shale formations; sane have been studied almost at the beginning of rotary drilling. A little progress has been made since then, but much is still to be learned. No simple solution exists so far, but good drilling practices combined with good mud practices have been helpful. It is well known that shales in general have a strong affinity for water, and when water is adsorbed, the stability of the shale section of a well is reduced; cohesive strength will be re- bed, properties will be altered and shale will generally expand and/or crumble. The amount of water a shale will adsorb depends on the characteristics of its clay mineral and the ionic concentration of surrounding fluid. Fluid loss and the resulting hydration of an unstable formation can lead to serious downhole problems. The rate of fluid loss depends greatly on the permeability of the deposited cake. In impermeable formations, however, there is no loss of fluid into the formation. As such, shale, because of its very low porosity and almost no permeability, could be termed impermeable. Yet there is strong evidence that shale surfaces are wetted by an invasion of fluid into the formation through microfractures and jointed planes. It is this wetting that causes the shales to swell and slough into the hole, resulting in problems such as increased mud volume and treating costs, bridges and fill-up, stuck pipe, overgauge hole, poor cement jobs and increased cement requirements, and logging and completion difficulties. Drilling operations have been defeated by severe shale problems, resulting in & hole being plugged and abandoned after several weeks of drilling. Shales are we1l known as being difficult and costly to drill. Numerous fishing jobs have plagued drilling operations, and the cost of reaming tight holes ad weighting up the drilling mud to control shale sloughing has been high. The concept of controlling water loss to stop sloughing was introduced in the 1920's. It was known-then that calcium in fresh water re- strains clay £ran excessive swelling (theory of osmosis) . This reinforced the belief that calcium-based muds such as lime, gypsum and calcium chloride should minimize hydration of clays. Lime muds were then introduced for this purpose and others in the 1930-1940 decade1. They were used for shale inhibition but were not effective. In an attempt to improve them, a special calcium-type mud beam popular during the 1950's. Hole enlargement continued and they too were considered ineffective. During & early 1960qs, papers were published claiming that shale swelling could be prevented by using high concentrations of chemical thinner called Chrane Lignosulfonate1. This began the era of highly treated lignosulfote muds. Unfortunately, the hole enlargement problem continued. During & mid-1960’s, oil based muds were used for the purpose of inhibition of shales. Not until 1966 did Mondshine2 suggest that the water phase be saturated with calcium chloride. The cost of using oil based mud was considered by most operators not worth the additional hole stability. The potassium chloride-polymer muds were also used in many fields in the United States and western Canada. The use of these muds has been claimed to prevent clay swelling by inhibition and encapsulation in sane areas of western Canada. They are also expensive and their ueneral application is still experimental. To reduce filtrate loss when sloughing shale becomes a problem during drilling is a reasonable practice. This study will investigate the use of cationic organic additives to mud to form a protective barrier between the shale - which are primarily composed of clays - and water that will keep water from entering the microfractures and will also prevent them from hydrating. In so doing, an understanding of modern clay chemistry and & dispersive properties of cationic additives in water based systems will be our most valuable tools. This study will also present and solve a mathematical model for simulating cation transport phenomena in fractured shales using the Framdlich equilibrium adsorption isotherm. The studies will analyze wellbore failure and stability as affected by cationic organic material.

Jan 1, 1986

By Waclaw Trutwin

The idea of automatic or remote control of the mine ventilation process generally, and methane concentration particularly, attracts the attention of mining engineers more and more. The advantages of introducing mine ventilation control systems are breaking traditional reluctance. The change of attitude is not only because of the requirements of modern exploitation technology, but it is also due to the recent progress in development and successful introduction of reliable monitoring systems and actuators in the form of controlled ventilators and doors [1]; [2], [3], [4], [5], [6]. Many 'years of theoretical and experimental studies of the dynamics of mine ventilation processes created the needed base for a proper design of an automatic control system [7],[8],[9], [10]. From these studies must, however, be drawn a fundamental conclusion, which may be regarded as the motto of this paper: An automatic control system for mine ventilation ill-conditioned or improperly designed is capable of creating hazard situations in response to random disturbances, much more, severe in consequence than a traditional ventilation system without any automatic or remote control! This statement is easy to prove if the dynamic properties of the ventilation process are taken into consideration. The ventilation process, as a matter of fact, is described by non-linear equations, and it must be expected that the process has more than one state of equilibrium. In other words, in the ventilation process may exist not only one but also more than one steady-states of flow, of which some are stable and others unstable. In certain circumstances, there may be no steady-state at all, and the process will oscillate [8], [11] , [12] . The state of flow in a network tends towards a steady-state and the actual steady-state established will depend on the initial conditions or disturbances in flow (fire,. etc.), which steady-state from the total number that will be . We frequently observe jumps from one steady-state to another. Disturbances in flow conditions which may cause such transitions are events of random character, occurring very rarely. Concluding, it must be stressed that there has to be a control system adjusted to the ventilation process in order to avoid situations mentioned above. There is only one alternative available and suitable for examination or study of the dynamics of a given mine ventilation problem: either by continuous monitoring of the real process, or numerical simulation of the process using a mathematical model. The advantages of the second method are obvious. This method allows consideration of every possible case very quickly and cheaply in relation to the first method. The aim of the paper is to show again that the simulation of the mine ventilation process and particularly a methane concentration process, separately or combined together with a control system, are real possibilities. A simulation method requires precise specification of the problem under consideration. For example, if we intend to examine a methane-concentration control system, the following items have to be specified: - expected target function of the control system. - structure of the control system. - mathematical model of control system, including sensor system, data preparation system, controllers, decision routine, regulators, etc. - structure of mine ventilation network. - mathematical model of ventilation process, including air flow and methane concentration processes. - pattern of disturbances which may occur in the controlled process as well as initial conditions on a 'start-up' of the system. Using typical computer programs for numerical solution of equations in the mathematical model of the problem involved, we are able, within the adequacy of the model, to simulate every case specified by the disturbances and initial conditions. As a result of simulation, it is expected that the following parameters could be defined: - transient flow in the network. - transient state of methane concentration in working areas. - stability of flow and methane concent¬ration. - stability of the control system. - range of control. - efficiency of control, etc. It is obvious that simulation methods readily allow for modifications to existing systems such that desired results will be obtained. Also optimisation problems could be solved by use of the simulation methods. In order to illustrate these general thoughts, a brief presentation of a mathematical model of methane concentration and

Jan 1, 1980

By William A. Rhoades

INTRODUCTION Ever since miners were faced with the task of moving ore, some form of scraper has been in use. At first, men and beasts of burden supplied the power to move the scraper, and later, machines were developed for this purpose. In the early 1900s, a few mining properties used small pneumatic single-drum winches to pull loaded scrapers to a raise or ore pocket, and the empty scraper then was dragged back to the muck pile by a miner, as shown in Fig. 1. Just prior to 1920, an improvement was introduced, using two single-drum hoists. As illustrated by Fig. 2, the second hoist was used to return the empty scraper to the muck pile. However, this arrangement still re¬quired two men, one operating each hoist. The next developmental step was to eliminate the second man by locating the hoists side-by-side and using one man to control both hoists. As illustrated by Fig. 3, this involved the use of a tail rope over a sheave to pull the scraper back. The greatest progress in development and the great¬est increase in the use of slusher haulers occurred between 1920 and 1930. In 1921, the Sullivan Machinery Co. designed and built the first two-drum scraper oper¬ating on the principle illustrated by Fig. 4. It was powered by a 4.5-kW (6-hp) Turbinair(r) motor and would pull a 450-kg (1000-1b) load at 0.61 m/s (120 fpm). In 1922, this unit was shipped to the Verona Mining Co. of Caspian, MI, and it experienced immediate success in the Lake Superior iron ore district. Since the two-drum slusher was much less expensive and more efficient than hand mucking, the Lake Superior mines were saved from financial disaster when iron-ore prices fell 25% between 1923 and 1925. Immediately there¬after, a demand developed for slushers that would oper¬ate with electric power, which was considerably cheaper than compressed-air power. In 1923, the Sullivan Machinery Co. responded with the first electric-powered double-drum hoist. During subsequent years, design improvements in¬cluded separate tail-drum gearing to increase the tail¬drum speed, as welt as a number of safety features such as rope guides. One result of these improvements and their utilization by the Michigan iron mines was an in¬crease of 100% in the tons of ore per miner per day in those mines between 1924 and 1929. Since the two-drum slusher was capable only of straight-line mucking, it was not a practical machine for use in open stopes. In 1929, the Sullivan Machinery Co. introduced the first three-drum slusher. As illustrated by Fig. 5, two tail drums and one hauling drum were provided. A tail sheave could be placed at each side of the stope, and the ore then could be loaded and hauled to a central point from the entire width of the stope. During the 1930s, progressively larger slushers were demanded. By 1940, two- and three-drum units were available with motor power as high as 45 kW (60 hp), and slusher power continued to increase after 1940. Be¬tween 1951 and 1952, Joy Manufacturing Co. designed a 112-kW (150-hp) two-drum Blusher for the Climax Molydenum Co. Although there has been no demand for a slusher more powerful than this, 150- to 225-kW (200- to 300-hp) slushers are quite feasible at the pres¬ent time. During the last 30 years, many slusher improvements have been made to the operating life, operational safety, and ease of operation and maintenance. Increased tail-drum speeds have decreased overall scraping times. Rope guards, totally enclosed drums, and operator shields have reduced the hazards of injury due to wire¬rope breakage. Improvements in lubrication have made the slushers relatively maintenance-free, with long operating lives. The introduction of spring-actuated drag brakes prevented uncontrolled unreeling of dis¬engaged drums, allowing the development of practical remote-control slusher operation. Remote control now is available in a choice of all-air, all-electric, or air¬electric slushers. APPLICATIONS Quite simply, slushers are used to load and transport material (ore), generally over a short distance of from

Jan 1, 1982

By H. Lyn Bourne

Daily everyone depends on the great variety of glass products, so much so that glass is often taken for granted. In fact most people do not realize how versatile glass has become. Consider the various uses and then try to imagine a day in which we are not influenced by glass. Common uses include container ware, table ware, window glass, lead crystal, automobile glass, and fiber glass. Several less common, but important, uses include laboratory ware, pharmaceutical, TV bulbs, light bulbs, glass ceramics, optical glass, fiber optics, and laser glass. Corning, Inc., a leader in specialty products, uses nearly 1 000 different compositions to manufacture about 60 000 different products (Edwards and Copley, 1977). Glass is such a complex product that-definitions vary and exceptions can be found for most definitions. Glass is an inorganic amorphous (non-crystalline) solid. Most glasses are produced by melting of a mixture of oxide raw materials, and then cooled to room temperature. Soda-lime-silica composition.s account for about 90% of all glasses melted (Anon, 1973). The properties of the glass product come mainly from its chemical composition. All of the different glasses require melting a combination of raw materials and forming the molten material into the desired shape. Both the melting and the forming processes use sophisticated technology and these technologies require experts to manage these production systems. The manufacturing process is continuous and takes place in tonnage quantities, so adjustments in the batch to achieve the desired finished product requires a great deal of expertise. Raw materials are fed to the batch mixing area in very large quantities (tons in most cases). As a result, impurities in the range of 0.1% result in addition of that impurity within the molten glass in kilogram amounts. More than twenty different industrial minerals are consumed in the manufacture of various kinds of glass (O'Driscoll, 1990). This chapter describes the major and minor ingredients of the various glass batches. It discusses the roles of the various oxides in the glass batch and most importantly considers the mineral raw materials which supply the glass industry. Each of the raw materials is described in detail in other chapters so the geology and mineralogy sections are kept brief here. Container glass, by far, accounts for the most production; followed by flat glass, fiber glass, and specialty glass of which table ware accounts for the greatest tonnage. [Table 1] shows the general production data for 1987 through 1990. Statistics for many of the uses do not appear because production volumes are small compared to the major uses. The glass industry is organized in four categories: containers, flat glass, fiber glass and specialty glass. The US Department of Commerce, Bureau of Census, publishes production data about the glass industry in three different categories: 1) glass containers, 2) consumer, scientific, technical and industrial glassware, and 3) flat glass. The Bureau has very complete statistics about the glass industry in these three categories but they report production data in different units according to industry standards. Therefore, [Table 1] gives the production data in dissimilar units. The production of most glass articles follows similar steps. The raw materials are mixed and the resulting batch is fed into the furnace. In soda-lime-silica glasses melting begins between 600 and 900°C. At these temperatures CO, and other gasses are released which create bubbles in the molten glass. To remove the bubbles and insure complete melting the temperature is raised to between 1 500 and 1 600°C. This is the melting-refining stage during which the refining agents in the glass batch serve to aid in the release of gas bubbles, homogenize the melt, and prevent the formation of scum on the surface of the molten liquid. At the conclusion of the melting-refining stage the glass is too fluid for working and the melt is cooled to about 1 100°C to attain the proper viscosity for working and forming to begin. After the glass article has been made, it must undergo annealing (slowly and uniformly reheated and cooled) to remove thermal stresses that were created during the forming process.

Jan 1, 1994

By G. Kraemer, J. A. Le Gac, P. R. ZETTWOOG

I. INTRODUCTION During the course of our activities in assisting mining companies, we have had access to the monitoring results for personnel from a large number of uranium mines. The results differ greatly from one mine to another. One of our objectives has been to discover the origin of these differences. It is evident that they are largely due to variations in the geological context, to mining methods, and to the organization of the work ; but we also found that the rigor with which measures are implemented to prevent the personnel from being exposed to radiations is also a cause. In order to advise mining operators effectively, we have asked ourselves the following question Given the dosimetric results of a mining site, how can it be known - if occupational exposure has been reduced to the lowest possible level (dose optimization principle recommended by the ICRP) ; - if the exposures were justified, which would not necessarily be the case following a questionable choice of the mining method, for instance, or a lack of efficiency in the maintenance and repairs of ventilation apparatus or even an excess of radiation protection. It is necessary to establish some criteria, while taking into consideration the specific conditions for each mine, in order to determine whether a mining company is adequately implementing radiation protection procedures. This need led us to attempt a modelization of the occupational exposures of uranium miners ; the preliminary phase is presented here. Although voluntarily still very basic, this model makes it possible to demonstrate the role of certain dynamic or passive variables. Moreover, the model presents the concept of specific irradiation of a mining site equal to the collective dose received per ton of uranium metal supplied to the uranium mill. The specific irradiation can therefore be used to indicate the effectiveness of radiation protection procedures at a given mining site. This model can be used for - previsional exposure studies based on the use of data gathered at each site, making it possible to compare various work methods and to determine prevention means ; - qualification of "radiation protection" procedures at a mining site ; - detection of unjustified exposures ; - research of ways to reduce inevitable exposures. 2. DETRIMENT TO BE ASSOCIATED WITH A GIVEN EXPOSURE DISTRIBUTION Respecting exposure limits ensures that delayed stochastic effects are the only health risks to workers. If the hypothesis of linearity in the dose-effect relationship is applied, the detriment to a group of workers is proportional to the sum of exposure for all the workers, or the collective exposure usually expressed as man x rem. On this point the linearity of effects hypothesis makes it possible to consider that two different standard deviation distributions having the same mean value are equivalent. It would be possible, then, to reduce the standard deviation by changing the distribution of workers at the various working places without changing the mean exposure value. In other words, without taking into consideration the practical implications, individual limits can be respected by rotating the personnel or by reducing the individual working period which would be compensated by increasing the number of workers. The collective dose would remain the same, which would justify the use of the concept of collective exposure. 3. "RADIATION PROTECTION" QUALITY INDEX AT VARIOUS MINING SITES Each mining site is responsible for its own collective dose, but it is also responsible for producing a certain amount of uranium metal in the ore sent to the uranium mill. These two quantities must be brought together in order to establish the "radiation protection" quality index of a mining site. This quality index, called-specific irradiation (Ir) is expressed in rem.ton -1 for external irradiation, in mJoule.ton-1 for inhaled [a]-energy, and in Ci.ton -1 for inhaled radon *. The specific irradiation represents the health hazard which must be associated to the extraction of one ton of uranium metal. Figures la and lb present a system of coordinates for which the axis of the abscisses is graduated in tons of U metal and the axis of the ordinates is graduated in units of collective exposures. In the EEC, the regulation is based on the measurement of the amounts of radon 222 inhaled.

Jan 1, 1981

By W. R. Dengler

INTRODUCTION Shaft sinking represents a relatively small proportion of the activities involved in mine development. Because of this limited utilization, shaft sinking has been one of the most difficult functions to mechanize. Although shaft machines (muckers) were introduced more than 30 years ago, the first machines were low-capacity units, and machines of this type did not gain wide acceptance by the mining industry until the late 1950s. However, the increasing depth of underground mines and the es¬calating costs of labor have increased the benefits obtain¬able from high-speed shaft sinking. The result has been a growing need for high-capacity shaft machines. Most of the machines available at the present time have been developed by modifying existing equipment to increase the payload capacity and performance. AVAILABLE SHAFT MACHINES In general, shaft machines may be classified accord¬ing to the types of digging mechanisms and the mecha¬nisms of operation. The available digging mechanisms include clamshells, cactus grabs, front-end buckets, and backhoes. The mechanisms of operation can be classified as either pneumatic or hydraulic. Selection Criteria For efficient operation, the primary consideration in selecting a shaft machine is its capacity. The minimum capacity of the machine should match the capacity of the hoisting system at the ultimate depth of the shaft. In many cases, two machines may be needed to obtain a sufficient mucking capacity. The shaft dimensions and configuration often dictate the type of machine that may be utilized; for example, an overshot loader does not work well in a circular shaft that has a diameter of less than 5.5 m (18 ft). Similarly, a single Cryderman mucker is not suitable for a large-diameter shaft, nor is it applicable where very high sinking buckets are to be used. Other factors for consideration include the degree of material fragmentation, the abrasiveness of the rock, and the amount of ground water that is inter¬sected. Machine Descriptions Riddell Mucker: The Riddell mucker, illustrated in Fig. 1, was first introduced in 1948. This is one of the most common shaft machines in the medium-capacity range. The Riddell mucker consists of a hoisting and traveling mechanism that operates a clamshell suspended on cables. The early machines operated with a 0.3 m3 (10 cu ft) clamshell and had an hourly capacity of approximately 27 t (30 st). This type of shaft machine has gained wide acceptance throughout North America. Essentially, the Riddell mucker consists of standard components, such as the pneumatic tugger hoists that operate the clamshell. The carriage is suspended on rails that also serve as the blasting set, located immediately below the shaft timbers. To effectively position the clam¬shell in the shaft, the operator controls the carriage travel by means of wheels that are driven by a pneu¬matic motor. The Riddell mucker most commonly is used in rec¬tangular timbered shafts. However, some units have been installed with circular rails, allowing them to work in circular concrete-lined shafts. A popular practice with many Riddell muckers is to increase the capacity through the use of larger clamshells, often ranging up to 0.76 m3 (1.0 cu yd), in conjunction with more powerful tugger hoists. In some large rectangular shafts, two Riddell units have been used side by side to achieve a sufficient capacity. Cryderman Mucker: The Cryderman mucker, illus¬trated in Fig. 2, is another common shaft machine in the low- to medium-capacity range. Developed in the early 1950s, it also has gained wide acceptance throughout North America. Equipped with a 0.3 m3 (10 cu ft) bucket, the Cryderman mucker has an hourly capacity of approximately 23 t (25 st). The Cryderman mucker is operated entirely by means of pneumatic cylinders and a pneumatic telescopic boom. The boom is suspended from a sectional frame¬work that is usually located at one end of the shaft. The framework or headframe can be suspended from an independent hoisting system located either on the sur¬face or in the shaft. In large shafts of either rectangular

Jan 1, 1982

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

By Gunter Zimmermeyer, Hartmut Eicker

Radon in the Environment Radon, as a natural nobel gas, can be detected nearly everywhere in the environment as a decay product of ubiquitous uranium. As it is emanated from soil and rocks measurable concentrations have been found at the surface of soils and in even higher concentrations in enclosed spaces like, for example, mines and buildings. While above soil surface activities caused by radon have been found in an order of magnitude of up to 1 pCi/l (Weigel, F. 1978), concentrations in enclosed spaces and mines are higher because of the lack of atmospheric circulation. Beside air circulation the relevant figure depends on the Ra226-concentration in the surrounding rocks or building material, as well as on emanation coefficient and the diffusion coefficient. While representative Rn222concentrations in well ventilated buildings are reported to be in an order of magnitude of 1 pCi/l maximum values up to one order of magnitude higher have been found in badly ventilated brick buildings (Ettenhuber, E., Lehmann, R., Clajus, P., 1978) (Aitken, J.H., et al., 1977). Just now it was stated that the reduced air circulation due to German legal regulations on energy conservation will increase radon exposure of the public considerably (Jacobi, W., 1979). Radon in Mines Radon exposure of workers is, of course, a matter of concern in uranium ore mines where relatively high concentrations of the uranium to be mined are present. Measures to protect workers' health have been implemented, based on experience on dose-effect relationship. They serve to meet exposure standards by limiting inhalation of radioactive particles, in reducing radon concentrations or in limiting working hours. Both improved measuring devices and capacity as well as the lower discrimination threshold enable to measure radon concentrations in other mines, e.g. in coal mines. It is known that radioactivity in coal is small compared with that in other minerals and even soil, rocks. Nevertheless, radioactive elements were identified in coal and so the question was whether the concentrations of radon in coal mines might be a subject of concern. The problems encountered when measuring radon in coal mines are described below, as the measuring device has to be flame proofed which is an important additional requirement. Measured radon concentrations in British coal mines have already been published (Duggan, M.J., Howell, D.M., Soilleux, P.J., 1968 (Dungey, C.J., Hore, J., Walter, M.D., 1978). The authors found concentrations of up to 14 pCi/l in Cornish mines. In most cases the values were in the order of 2 pCi/l. These results were consistent with measurements reported from U.S.-coal mines (Lucas, H.F., Gabrush, A.F., 1966). Such concentrations of radon were not considered to represent a hazard for British miners (Ogden, T.L., 1974). In Germany, too, first measurements have been carried out in five coal mines in the Saarland in the 60's. Air samples were taken at different places in the coal mines, dried, fed to an ionisation cell and measured by a device including reference cells. Samples taken at ventilated places showed radon concentrations consistent with the lower British results. They all kept within the standards of the first German regulation on protection against radiation. Measuring the radon daughters was renounced because of the relatively low radon concentrations and the requirements for flame proofness in coal mines. Moreover, it can be ascertained that because of the effective ventilation the disequilibrium factor between the decay products and the radon concentration remains far below the value of one (Muth, H., 1978) (Keller, G.). In 1979 the committee on mine safety and health protection in coal and other mines of the EEC proposed to have measured and evaluated radon concentrations in European coal mines to find out whether they complied with international standards. Great Britain and Germany agreed to this proposal and by commissioning such measurements to scientific institutes complied with the request to harmonize the methods used. In the Federal Republic of Germany, e.g. Westfälische Berggewerkschaftskasse (WBK) and Staatliches Materialprüfungsamt; Dortmund (MPA) were requested to carry out the measurements in coal mines of the Ruhr coalfield whereas Saarberg Interplan was responsible for the Saar coalfield. The WBK measurements are reported in later paragraphs.

Jan 1, 1981

By N. Piret, B. Shoukry, S. Buntenbach

Traditional or artisanal goldmining, also known as small scale goldmining, has a strong and probably a negative environmental impact. The processing methods applied are very frequently a source of severe pollution due to the emissions of mercury by the extraction of gold by means of amalgamation as well as the emissions of cyanide through cyanide leaching of gold bearing ores. The emissions find their way into the environment and contaminate soils, sediments, water and atmosphere. Abnormal concentrations of mercury and cyanides in waterways are known to occur year after year destroying irreplaceable regions of the world. Mercury and cyanide compounds are highly toxic and may directly create permanent damage to the whole ecosystem. Existing methods for recycling of mercury and for decontamination of mercury and cyanide contaminated tailings are not customary applied in small scale mining and are ineffective as well. Based on investigations of traditional and small size goldmining, this paper presents: -processing methods of gold and discarded tailings under consideration of environmental protection; -figures on actual situation; -recommendations for equipment; -some decontamination methods for mercury and residual cyanide. Mineral Processing methods in traditional gold mining Gold is usually existing in its ores as the metal alloyed with metallic silver and perhaps copper. The element may occur in the form of: -native gold -inclusions also of microns or submicroscopic size metal sulfides (auriferous) such as pyrite, pyrrhotite, stibnite, arsenopyrite and galena -combined as telluride or sulphotelluride. The separation process selected depends on whether the gold can be freed from its unfavorable associations (e.g. gangue) at a sufficiently coarse grain-size, or whether it is carried in a heavy sulfide which can be freed similarly. The usual practice is to concentrate the goldbearing mineral at a relatively coarse grain-size and to regrind the ore if necessary. The gold content is concentrated by secondary or tertiary gravital methods or is extracted by chemical methods (amalgamation, cyanidation etc.) Gold, even when of fine grain-size, settle readily due to its high specific gravity from pulps in which the main gangue mineral is quartz or silicates. Amalgamation is the process of separating gold and silver from their associated minerals by binding (entrapping) them into a mixture with mercury. The cyanide process is applied to separate gold or gold-bearing compounds by dissolution from the finely ground ore (CIP, CIL, RIP), or as heap leaching. The dissolved gold is separated from the solids and the metal-rich or pregnant solution is then treated to recover its gold. Gold is also recovered by flotation methods. This process is widely used in treating base metal ores and in separating various sulfide components of ores, as well as in removing the barren gangue. The gold usually associates with a specific product in a sequence of flotation operations and is recovered subsequently in the smelting of the sulfide concentrates and refining of the metallic products, or by cyanidation of the roasted concentrates. Froth-flotation can be applied to separate gold and sulfide minerals from a finely ground pulp. The Amalgamation Process Amalgamation is the main method for the recovery of gold in traditional mining and is applied for the extraction of gold from placers as well as primary ores. The mineral technology used depends on the nature of ore deposits. In winning gold from solid ore, the matrix of minerals and rocks must be crushed and ground to sufficient fineness to liberate the gold. The liberated gold could be treated similar as free gold from placers. Gold is mainly separated from the valueless gangue (barren rock) by utilizing the difference between the density of the impure native metal (density about 16-19) and the gangue (density about 2.5). In simple operations the material is carried by a stream of water down a sluice generally equipped with small transverse barriers (riffles) against which the gold collects. The riffled sluice is the principal device used by artisanal gold miners. Nowadays, spirals as well as centrifuges, such as Knelson separator or Falcon separator, are occasionally applied for gold recovery. Gold may also be recovered from the pulp, by passing it over corduroycovered tables that catch the heavier particles - a method maybe as ancient as gold mining itself. In history, sheep skins were used to catch gold particles in this manner. Furtheron, gravity separation of gold is practiced on jigs, hydraulic traps, shaking tables and

Jan 1, 1995

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998