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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 R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981

By Steve Karl

President Reagan may be "a nice guy," but he is "misinformed, misdirected, and misadvised," when the subject is the state of the US copper industry, according to Sen. Dennis DeConcini (D-AZ). DeConcini took the opportunity as keynote speaker at the Arizona Conference AIME in Tucson to fire a few salvos at the Reagan Administration's industrial policies. "American copper used to stand above the rest of the world," he said. Now 21,000 copper workers, about half of the total, are out of work due to less expensive foreign imports. "Those 21,000 are real people, not statistics," he said. US production has been cut to one-third of its capacity, he said. And the Administration shows no signs of changing its position to favor US copper protection. "Third world copper towns are booming," he continued, "while ours are dying." Regardless of profits and despite oversupply, Chile continues to produce, he said. And, while US mines continue to close, "the International Monetary Fund (IMF) is handing more than $1 billion to six copper producing countries." President Reagan wanted $8.6 billion from the IMF. "I'm damn mad about it," DeConcini said. "For the life of me, I can't understand how this Administration can stand by while this industry is brought to its knees." Last year, the International Trade Commission ruled that imports were injuring domestic copper and recommended relief. The President, DeConcini said, vetoed those recommendations. DeConcini softened his tough talk a bit saying the President's image makes it difficult for people to not like him or stand up to him. "How can anyone stand up to President Reagan?" he asked. "He's such a nice guy. But it's time someone did. He's just misinformed, misdirected, and misadvised. We must take real action and we must have a president who understands this." DeConcini said he has introduced legislation aimed at helping domestic copper. It would limit copper imports to 385 kt/a (425,000 stpy). Imports now stand at about 635 kt/a (700,000 stpy). The bill would also impose a $0.33/kg ($0.15-per lb) duty on foreign copper. DeConcini called the duty a sort of "environmental equalizer" because that is the amount domestic producers must spend on pollution control devices. Foreign competitors do not have such controls, he said. "I face people who are damn mad that this country is being pushed around," he concluded. "It's time we stand up and say we can be competitive. If they (foreign countries) put an import duty on our stuff, we will do the same. It's time this country stopped being the nice guy." As if to underscore domestic copper's desperate situation described by the Senator, Duval Corp. announced about the same time as the meeting that it has nearly closed its eastside office in Tucson. Staff has been reduced from 120 to four. Spokesman Dean Lynch said the four will consist of President A. Everett Smith, a secretary, a person in environmental affairs, and another in purchasing. Duval is also selling an office and a laboratory in Tucson. Pennzoil Co., Duval's parent, has been trying to sell the company for more than a year. It began dismantling Duval in November 1984. Pennzoil took over its subsidiary's profitable sulfur operation in Texas, sold the New Mexico potash facility, and spun off gold interests in Nevada, forming Battle Mountain Gold. Northwest Mining Association - Spokane Rock Jenkins, Associate Editor The true role of minerals needs to be realized by both the policy makers and the people of the US, according to Robert Dale Wilson, director of the Office of Strategic Resources, US Commerce Department. In addition, a re-thinking of the theory of free trade and competitive advantage is necessary. Wilson made his remarks in December at the opening luncheon of the 91st Annual Convention of the Northwest Mining, Association in Spokane, WA. At a later press conference, Wilson said one of the mining industry's main problems is that its presence in Washington has been reduced in the past few years. Part of this can be seen by events within the American Mining Congress (AMC), he said. "The problem with AMC," Wilson said, "is that in 1981, when Reagan came in, no problems were seen for mining and a lot of their (AMC's) lobbyists were let go." He

Jan 1, 1986

By Kelly F. Bailey

Introduction On June 20, 1986, the Occupational Safety and Health Administration (OSHA) published revised asbestos exposure standards for general industry and construction. The standards reflect OSHA's attempt to adequately control workplace exposures to minerals it considers carcinogenic - minerals capable of causing or contributing to cancer. These standards specifically identify asbestos as: chrysotile, an asbestiform serpentine mineral; and the amphibole minerals amosite, crocidolite, tremolite asbestos, actinolite asbestos, and anthophyllite asbestos. Each of these has a more common nonasbestos mineral analog that exists in nature in a crystalline, blocky shape rather than the hair-like or fibrous shape of asbestos. The mineralogical names for three of these nonasbestos minerals are unique: antigorite for chrysotile, cummingtonite-grunerite for amosite, and riebeckite for crocidolite. The other three nonasbestos analogs do not have unique mineralogical names. They are simply designated as actinolite, tremolite, and anthophyllite without the word asbestos following their names. The 1986 OSHA standards not only cover exposure to the six asbestos minerals, they also cover specifically the nonasbestos forms of actinolite, tremolite, and anthophyllite (AT&A). The new standards regulate these minerals exactly like asbestos (OSHA, 1986). The construction aggregate industry views this as a major problem because these nonasbestos minerals are common amphibole rock-forming minerals in the earth's crust. They exist in small quantities over large areas of the United States (Kuryvial et al., 1974). These minerals, unlike asbestos, are not mined for a specific commercial purpose. They are unavoidable components in much of the aggregate used for construction throughout the US. They are also common in the gangue material of metallic ores. There are areas of the US where amphibole-bearing bed¬rock is common. Not every rock mass in these areas contain amphiboles, however. It does mean, though, that amphiboles are physically compatible with many of the rocks in those areas. And given the correct geochemical conditions, they will be present primarily in the nonasbestiform variety. In addition, these amphiboles will probably exist in the natural drainage system, sand and gravel deposits, stream sediments, lake shores, valley basins, or ordinary beach sand within these areas. There has been little quantification of nonasbestiform AT&A in dusts and soils in the US. This is not surprising since these nonasbestiform minerals are not commercially valuable. However, an example of the pervasive nature of these minerals can be found in a 1981 Geological Society of America publication where about 0.7% tremolite-actinolite was found in the desert dust in and around Tempe, AZ (Pewe, 1981). Since OSHA standards treat these common nonasbestos minerals as carcinogens in the same way as asbestos, large natural areas in the US are implicitly being labeled as hazardous by OSHA. When a substance is identified as a carcinogen, another OSHA standard comes into play, the Hazard Communication standard. There are also right-to-know laws in 9 states that essentially duplicate this federal standard. These standards require that a product containing 0.1% or more of an OSHA-designated carcinogen be labeled as such (OSHA, 1983). This means that much of the stone and sand gravel products occurring naturally and mined in the US could be labeled a carcinogen when, in fact, they are not. The National Stone Association (NSA) and the domestic construction and mining industries believe that OSHA has seriously erred. The NSA has studied the health, mineralogical, technical, economic, and legal basis for OSHA's action. These studies concluded that there is no justification for the agency regulating nonasbestos minerals as if they were asbestos. Health issues The preamble to OSHA's 1986 asbestos standard states that evidence for asbestos-like health effects from exposure to nonasbestiform varieties of AT&A is inconclusive (OSHA, 1986). The fact is, not only are the data inconclusive, they are nonexistent. During 1986-1987, NSA's occupational health and epidemiology consultant, Environmental Health Associates (EHA), reviewed all available health studies related to AT&A. EHA found evidence that malignancies in both experimental animals and humans are associated with the asbestos forms of these minerals. No experimental or epidemiological evidence was found that indicated such pathogenic effects occur from exposure to nonasbestiform varieties of these minerals. There are relatively few scientific studies of the health effects of exposure to nonasbestiform varieties of AT&A. In three different animal studies, exposure to either nonasbestiform tremolite or actinolite did not result in pulmonary fibrosis on in excess tu-

Jan 11, 1988

By Guangming Zhao, Nianjie Ma, Demao Guo, Denghong Chen, Yingming Li

Based on the mechanical essence that large-scale plastic failure zone appears in all or part of surrounding rock in soft rock roadway, the numerical simulation method is used to study the rectangular roadway in layered rock strata. It is clarified necessary conditions must be met for soft rock: firstly, the strength condition is that the maximum confining pressure is greater than the uniaxial compressive strength of rock strata. Secondly, the stress environment condition is that the ratio of maximum confining pressure to minimum confining pressure is greater than 3. Thirdly, the angle condition is that The direction of principal stress action enables the plastic zone of weak rock layers to fully develop. At the same time, the quantitative description method of soft rock is given, and the soft rock roadway is redefined. Soft rock roadway refers to the roadway that meets the strength conditions, stress environment conditions, and rock structure angle conditions under certain surrounding rock conditions and in-situ stress environment conditions. After the excavation of the roadway, a large-scale plastic failure can be formed, that is, a butterfly-shaped plastic zone is formed, and the conventional support cannot be adapted. It is difficult to support in engineering. It provides a theoretical basis and engineering analysis method for the identification of soft rock roadway, and the research results have engineering value Soft rock tunnel engineering in coal mines constitutes a vital aspect of soft rock engineering. This field broadly encompasses rock engi- neering concerning large plastic deformations, e.g., soft rock slope engineering and soft rock tunnel engineering. The intricate geological conditions encountered in soft rock tunnel engineering present a significant challenge to support, which has harmed coal production in China. China leads global raw coal production with the annual output of 4.6 billion tons. Annual tunnel excavation supporting this production spans approximately 11,000 km, with over 10% of these tunnels classified as soft rock formations. Soft rock is commonly associated with soft rock tunnels due to their prevalence in engineering projects. However, reaching a consensus on the definition of soft rock has long been an enduring challenge for scholars and engineers. Numerous definitions have been proposed, includ- ing descriptive, index, and engineering definitions. For instance, the International Society for Rock Mechanics defines soft rock based on its uniaxial compressive strength σ ranging from 0.5 to 25 MPa. China's Engineering Rock Body Standards, established in 1994 (GB 50218-94), take a qualitative and quantitative approach to classifying rocks. Rocks are categorized as hard or soft based on criteria such as hammering sound, fragmentation, water immersion effects, and weath- ering degree. Additionally, the integrity of rock bodies is assessed across five categories intact, relatively intact, soft fractured, fractured, and extremely fractured. This classification considers factors like the number and spacing of structural planes, their combination, and the types of structures. Descriptive and index-based definitions fall under the category of geological soft rocks, providing a comprehensive geological perspective on the surface features or strength characteristics. However, these definitions have limitations in engineering practice, which leads to contradic- tions. For instance, rocks with uniaxial compressive strength less than 25 MPa may not exhibit soft rock characteristics if the tunnel is shal- low with low horizontal stress levels. Conversely, rocks with compressive strength exceeding 25 MPa at sufficient depth and high horizontal stress may exhibit soft rock characteristics. Definitions originating from engineering practice have emerged after realizing the inadequacy of discussing soft rocks without considering engineering. For instance, Dong's loose circle theory defines soft rocks as rocks with a loose circle thickness exceeding 1.5 m, which chal- lenges conventional supports. This intuitive definition, widely accepted by engineering professionals, emphasizes the difficulty in supporting tunnels due to extensive damage. However, various tunnel damage poses a challenge in relying solely on the loose circle thickness of tunnels for determining soft rocks. He introduced the concept of engineering soft rocks, which are defined as rock formations exhibiting significant plastic deformations under applied engineering force. Two fundamental mechanical properties of soft rocks are identified the critical softening load and critical soft- ening depth. Rocks below the critical softening load threshold are categorized as hard rocks, while those exceeding it exhibit substantial

Jun 25, 2024

By Elza Burch

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

Jan 8, 1987

Power cable costs are only a small part of total mining costs. So many mine operators consider power cable failure and resultant downtime as part of the cost of doing business. But, viewed in terms of lost production, these costs can be quite significant. Now one company, Cablec, seeks to cut cable costs by upgrading the polymer compounding process used to make cable insulating and semiconducting materials. Cablec is the leading manufacturer of electrical power cables in North America. And with about a third of the market, Cablec is the largest supplier of power cable to the mining industry in the United States. To improve its products, Cable has entered the polymer compounding business. In July, it began producing insulator and semiconductor polymer compounds at its plant in Indianapolis, IN. "This new facility provides a quantum leap over conventional compounding methods," said Harry C. Schell, Cablec's president and chief executive officer. "The Cablec polymers plant is producing a dramatically higher standard of polymer compounds that provide significantly higher levels of performance and improved life cycle costs for power cable." Cablec faces tough foreign competition in the wire and cable business. Competing on price alone is difficult, particularly when foreign producers are state subsidized. So Cablec feels the best way to compete is to establish new quality production standards. The company's new polymers plant is one way to do this. By increasing purity control and uniformity in polymer compounding, Cablec says its power cables will last longer and fail less often. A typical medium voltage cable consists of a conductor, conductor shield, insulation, insulation shield, metal shield, and jacket. The conductor shield and the insulation shield are conducting polymers. Contaminants and imperfections can occur within the insulation, at the conductor shield/insulation interface, or at the insulation shield/ insulation interface. Over time, these contaminants and imperfections can decrease the electrical strength of the cable or cause premature cable failure. The effort to minimize the number and size of any possible contaminants begins with pure polymer compounds mixed in a clean facility. However, most power cable manufacturers manually handle raw materials, use ethylene/propylene (EP) in bulk bales, and mix polymercompounds in open Banbury mixers. The quality and uniformity of polymer compounds is also impacted by temperature variations in the mixing process. This results in wide gradations of product consistency from batch to batch and ultimately contributes to power cable failure. Cablec says the improved polymer compounds from its state-of-the-art plant will be the purest and most consistent insulating and semiconducting materials available. The plant itself RCA spent $18 million to build Cablec's Indianapolis plant. RCA used the facility to mix specialty polymer compounds used to make video disks. RCA had two considerations in mind for the plant, cleanliness and uniformity of the compounds. However, when the video disk market failed to materialize, RCA sold the 46.5 dam 2 (50,000 sq ft) plant to Cablec for $3.1 million. Cablec invested an additional $3 million for modifications and increased production capabilities. Today's replacement cost for such a facility is estimated at $30 million. Cablec says the plant will set a new standard for performance and be economically difficult to duplicate anywhere. One of the essential elements of the plant's clean process environment is the air intake system. It filters contaminants greater than 2 um, less than one-fiftieth the current industry standard. All material handling and conveying areas in the facility are air-locked. This keeps out contaminants such as smoke, dust, and pollen. Banks of pneumatic pumps move polymer components through the system and continually filter the air. The plant also has a backup air intake system. No process downtime due to pump failure here. From the time raw material enters the plant, it is stored, transported, and processed in filtered air by an airtight stainless steel system. The stainless steel resists rust and corrosion. This further eliminates the danger of contamination from paint or rust particles in the conveyance network. A computer system allows a single operator in a central control room to monitor every aspect of the compounding process from air quality to line speed. The computer

Jan 12, 1988

By Dwane K. Johnson

INTRODUCTION The small mining company faces the dilemma of how to finance the development of its properties because it doesn't have the financial resources to pay for the development costs from its own funds and doesn't have the financial strength which will enable it to borrow the money needed for development. However, with a little imagination and the right set of circumstances, a financing can be arranged for the small mining company. This paper is an attempt to describe the fundamental steps on how such financing can be obtained --in other words--what is required to provide ad- equate capital at a low cost for developing and bringing a mining property into production. Innovative financing follows from this. The basic keys for building and maintaining a successful mining company are "the four M's": Management, Market, Mine, and Money. The writer believes their order of importance is as presented and each is a building block required in order to obtain the type of financing that best suits the borrower and lender. MANAGEMENT Management must be experienced, highly organized, realistic, able to take advantage of opportunities and fulfill promises. Mine financing requires a variety of skills and concentration on the part of management. Management should have a thorough understanding derived from experience in geology, construction, mining and metallurgy and environment, marketing, and financial matters. Management must realistically visualize what will be produced at what cost and truly evaluate the risks associated so the proper security can be correctly designed. Since risk cannot be eliminated, the objective of management is to identify the risk which will be present in a given venture and assess the level of that risk which will be acceptable to the firm. Generally the risk which is present is subject to little or no modification. After management identifies and measures the acceptable risk the firm will proceed in a way that it will be least exposed. Management must determine its long-term objectives and strategies within the context of a constantly changing world. This question must be addressed before examining the sorts of risk which affect the development and operation of a specific property. The compilation and interpretation of "hard data" by competent people is good to a point but the experience and instinct of seasoned individuals are the important factors in management choosing a long-term direction for the corporation. The attitude towards risk greatly affects the goals of the corporation. Major risks in a mining project can be placed in four categories: 1. Market - Price, demand, substitution. 2. Costs - Capital, operating, financial. 3. Regulation. 4. Taxation. The feasibility study for a project is the fundamental tool used in the management of risk. Management should employ an experienced team with an established and well-understood set of ground rules for the preparation and assembly of a feasibility study. The necessary degree of realism is built into all levels of the study and if the project is technically feasible, the company's hurdle rate is then used to discount the base-case project cash flow in order to determine its financial viability. Understanding the geological significance of the mineral deposit is critical. There should be strong interplay among the pure regional geologist, the detailed mine geologist, the mining engineer, and the metallurgical engineer. The collaboration among these different disciplines will yield a feasibility study that is more reliable. The essence of this work must be appreciated and understood by the executives planning the financing. It is vital that they thoroughly understand the risks and make the proper risk assessments to cope with our rapidly changing market environment. Projecting future revenue values is paramount and is subject to much estimation. One way to mitigate fluctuations and the risk of falling prices is by selling production for future delivery. MARKET Since mining is a worldwide industry it is mandatory that bankers be aware of what is occurring in the marketplace, both foreign and domestic. The marketing of the production is a very important facet when considering financing for the small

Jan 1, 1987

By Peter V. Avotins

During last year's AIME meeting in New Orleans, Deepak Malhotra and I were discussing some of the topics that would be presented in this Symposium. It seemed appropriate that along with papers on specific reagent and flocculant types, we should address research management aspects of new product development. This development is a long, arduous task that requires a significant technical and fiscal commitment by the chemical company and potential users. A flow chart depicting the research process is shown in Figure 1. The sequence of events are plotted on an undefined time line. Our experience has shown that it takes 3-5 years to complete a typical reagent development project. A new project starts with an idea of how to do something better. The ideas range from relatively simple product modifications to opportunities created by technological change. An example of this is the use of flocculants in tailings operations to close the water recycle loop. The change was brought about by the need to conserve water and comply with environmental statutes. The result was the development and wide-spread use of synthetic polymers to help in solid-liquid separation. More recently the ups and downs of oil prices have led to development of coal water slurries as an alternative to fuel oil. The slurries contain high levels of chemicals (up to 3%) to insure proper stability and burn characteristics. There are technical changes under way right now that may lead to ideas for other new products. One exciting source of new technology is the ability to manipulate genetic material of microbes to select for specific capabilities. This has evolved by leaps and bounds over the past 5 years. Biologically assisted leaching and removal of wastewater impurities are two possible end uses. I'm sure that there are many others. The ideas can originate anywhere and are often half-baked at first, but to lead to new products they must be refined to provide project definition and eventually a research plan. A chemical ,development project cannot be started without approval to commit the company's resources. The internal selling job is often very difficult. The company must see clear financial benefits in order to launch efforts that may not pay off for many years. It is critical to project the total market, profitability, eventual market share, and timing of introduction of the new product. There is much uncertainty in this and the research manager is often skating on thin ice to retain his credibility. Once we obtain support from upper management, the research team is assembled to start the project. The teams generally consist of organic chemists, research metallurgists, and technicians. They have access to analytical scientists and technicians for surface and analytical chemistry. The projects are often augmented by help from university staff and consultants. This is particularly true where there is a need to develop fundamental understanding of the technology in addition to synthesis of chemicals. The project now proceeds along two lines. First, there is an emphasis on understanding the prior art. Extensive literature and patent searches are undertaken to provide research leads and to keep from "reinventing the wheel" and infringing on someone else's patent. The scientists and technologists are qualified to conduct computer-assisted library searches. This process usually uncovers 80-90% of the relevant information. The remainder is very much harder to get and requires the help of professional library search staff with scientific backgrounds. Along with literature searching, the team begins synthesis of new compounds. These are screened by use of techniques such as microflotation or filtration in a Buchner funnel. The scientists are looking for the direction of response as they modify structural features . Inevitably, a large number of candidates reach the next stage, a laboratory bench test. At this point it is possible to set up a research plan and use project management techniques to monitor progress. The bench tests are conducted in larger equipment, usually requiring 100-1000 grams of test substrate. Typical laboratory flotation tests are run in Denver or Wemco cells.

Jan 1, 1986

By Carl Eschman

Productivity from large walking draglines is primarily dependent on operator skills. This machine may be in operation three shifts a day, 364 days a year, and its output is directly related to coal uncovered and mine profitability. Dragline operators must have highly developed manual skills and be knowledgeable in mine planning and working strategies. When using equipment costing more than $25 million, some formal training is usually required before an operator is allowed to assume complete con¬trol; however, dragline operators rarely receive any structured training before operating these giant excavators. A form of apprenticeship is usually followed where an operator candidate progresses from a groundsman to an oiler position. As an oiler, he is permitted to operate the dragline for short periods under supervision. After apprenticeship, the operator is considered sufficiently prepared to operate the largest, most powerful machine at the mine. The apprenticeship training method has obviously provided the surface mining industry with skilled dragline operators; however, conditions are arising that require a realistic and effective training tool that can be accessed by mining companies. New mines -either planned, under construction, or recently opened in the West-do not have access to a pool of experienced operators and oilers as do Midwest mines. As coal mining activities increase in both the West and Midwest, demand for trained dragline operators could be required in a short amount of time. Also, the more productive techniques along with sound basics of strip mining are sometimes lost in the informal "OJT" training method. Modern draglines are the pacemakers of the strip mine, and are simply too expensive to be used as a training device where lost productivity and susceptibility to damage can directly affect mine output. The Dragline Training System is a logical first step in formally educating or retraining operators. The program, started by the US Bureau of Mines and continued by McDonnell Douglas Electronics Co. under contract with the US Department of Energy, was installed and evaluated at DOE's Carbondale Mining Technology Center in Carterville, IL, last year. It is now being operated by Southern Illinois University at Carbondale. System Description The Dragline Training System addresses specific environments and work practices encountered during an actual mining operation at a midwestern US surface mine. This area was chosen because of its high number of strip mines using large walking draglines. Most draglines in the region are Bucyrus-Erie 1370, 1450, and 2570 models, so the dragline trainer was patterned after the company's 1370 machine. Operating and emergency controls are sufficiently standardized for most large walking draglines, and peculiar dynamics and responses from any specific dragline can be programmed into the computer system. The computer simply prompts the user to select the manufacturer and any peculiar response or rate changes needed. A 46-m3 (60-cu yd) bucket is simulated, but for closer simulation, various bucket configurations can be provided. Dragline Trainer The dragline trainer uses the TV-model simulation technique. A scaled model of the dragline is positioned in a model mine. A television camera is positioned at the operator's theoretical eyepoint, and the view captured is projected into a large screen in full scale. The screen is positioned in front of the operator seated in a full-size cab at Bucyrus-Erie controls. By manipulating the controls, the trainee can operate the model dragline and observe its reaction in the television display. In addition to housing dragline controls and consoles, the wooden, oversized cab contains the digital computer, terminal, video recorder and monitor, power switch box, air conditioner, and has enough room for the instructor and five student observers. The 50:1 scaled dragline model contains servo-controlled functions for hoist, drag, swing, delta swing, and longitudinal and lateral position. The delta swing provides bucket lag during swing and a realistic pendulum action when the swing is terminated. The over-responsive second order servo system is designed to provide hoist, drag, and swing rates exceeding present draglines. In all cases, position servos are used for better control and sta¬bility. The normal rate operation of an actual dragline is computed for the specific machine and presented to the servo amplifiers as iterative position commands increasing or decreasing

Jan 6, 1982

L.R. Bower In regard to the paper by J.B. Gwiazda, it makes a highly technical approach to show that the µ 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 µ 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. Supports to this design of both 450- and 280-t (496-and 309-st) rating have been successfully used in the United Kingdom for several years, negotiating many faulted areas without one single reported need for repair or maintenance. This includes supports left unattended during the year-long strike, proving the reliability of the system.

Jan 8, 1986

By Jon P. Thorson, James W. Babcock

The second Belt Symposium in Missoula, MT, Oct. 9-16, 1983, was designed by its conveners as a field conference on progress in the Proterozoic Belt Supergroup rocks. The symposium was sponsored by the University of Montana with corporate support by Anaconda Minerals. Asarco Inc., Exxon Minerals Co.. Hecla Mining Co., Molycorp Inc., Noranda Exploration Co., and Utah International Inc. Introduction The Belt Supergroup is a thick middle Proterozoic sedimentary sequence deposited in a basin along the edge of the North American craton. Subsequent geological events have modified the basin so its remaining rocks outcrop in Montana, Idaho, northeast Washington, southeast British Columbia, and southwest Alberta. The Canadian equivalent of the Belt is called the Purcell Supergroup. Several similar age basins occurred along the western edge of the Proterozoic North American craton. The sedimentary sequence in the Belt basin aggregates is more than 18 km (60,000 ft) in the thickest section, where the bottom is still unexposed and the top removed by erosion. The limits of the Belt basin are poorly known because of later sedimentation and tectonic disruption. Knowledge about the Belt terrane has been difficult and tedious to obtain because the rocks are generally monotonous and fine-grained. During the last decade, new 1:250,000 scale reconnaissance maps of most of the Proterozoic Belt terrane were published, open-filed or otherwise brought close to completion. These maps provided the opportunity and challenge to consider the stratigraphy and tectonics of the Belt Supergroup as a complete basin. This broader view has contributed greatly to the solution of numerous geological problems created by previous, more focused work. Along with the availability of better geological base, there have been new exploration developments. The 1973-1983 decade saw proof of the commercial viability of the Cu-Ag mineralization at the Troy mine. The conclusion was also accepted that the Sullivan mine was a sedimentary massive sulfide. So was the conclusion that the Cu-Co mineralization at the Blackbird mine at Cobalt, ID, was the result of similar submarine exhalative processes. And there were discoveries of another Cu-Ag deposit "larger and higher grade" than the Troy mine and of two new massive sulfide districts. The Belt-Purcell terrane has long intrigued exploration and mining geologists as the host of Pb-Zn-Ag veins of Idaho's Coeur d'Alene district and of the Sullivan mine at Kimberley, British Columbia, one of the largest Pb-Zn-Ag deposits in North America. In the mid-1960s, Bear Creek Mining Co.'s discoveries of the stratabound copper-silver occurrences in the Revett Formation brought new interest in the basin's geology. The recognition of the Sullivan ore body as a sediment-hosted massive sulfide has also generated minerals exploration activity in the US portion of the Belt basin. This activity increased again after the Australian discovery of the world class Olympic Dam deposit at Roxby Downs, South Australia. This occurred because of some suggested similarities in the regional geologic setting of the Belt basin with the Adelaide geosyncline, the host for the Olympic Dam deposit. Technical Sessions Technical sessions topics included Lower Belt and Ravalli Group stratigraphy, Middle Belt carbonate and Missoula Group stratigraphy, Belt-Purcell stratigraphy, structure and tectonics, strata-related mineral resources, geochronology and geophysics, geochemistry and isotope geology, and a summary of significant problems. Evening poster sessions dealt with the topics and papers presented each day. Since the first Belt Symposium, mapping of the stratigraphic units and their tectonic setting has enabled geologists to progress in conducting a basin analysis. Belt Symposium II revealed no concensus on many aspects of belt geology. However, basic stratigraphic studies have reached the point that models can be constructed. The Prichard Formation is equivalent to the Aldridge Formation, but detailed correlations were not previously available. Studies were made by Finch and Baldwin (Cominco American) and by Cressman (US Geological Survey). These studies established important correlations of key markers within the Aldridge Formation to units in the Prichard Formation. This will help eliminate geologic interpretation problems at the US-Canada frontier and make Belt basin reconstruction more accurate. Also, the Sullivan ore deposit was shown to be formed in the deeper distal part of the basin. This contrasts to the more proximal depositional position of the Prichard Formation. Hamilton

Jan 6, 1984

By Nenad Spoljaric

Greensand, greensand marl, and green earth are names given to sediments rich in the bluish green to greenish black mineral known as glauconite. The word glauconite is derived from the Greek word glaukos, meaning bluish green. The term "greensand" as a rock name for a glauconite-bearing sediment is more appropriate than "greensand marl," a term that has been doggedly perpetuated in the literature. Because of its potash and phosphate content, greensand was mined and marketed as a natural fertilizer and soil conditioner for more than 100 years. The advent of manufactured fertilizers with adjustable nutrient ratios led to a decline in the use of greensand in agriculture. The material has since been recognized as useful in water treatment. Unfortunately, despite large reserves and world- wide distribution, glauconite has not been utilized to any significant commercial extent because no major application has been found for a substance with its chemical composition and properties. This is probably due mostly to a paucity of research on its potential commercial uses. Extraction of potash received considerable attention during and just after World War I. Because of relatively high extraction costs and a generally low potash content (viz., less than 8%), glauconite lost its appeal as a source of this commodity. Historical Background Greensand was used as a fertilizer in New Jersey in the latter part of the 1700s. During the early 1800s its use became more common; applications of as much as 22.5 kg/m2 were sometimes made, although recommendations for agricultural use suggested 4.5 to 11 kg/m2 (Tedrow, 1957). Many crops, especially the forage type, were said to improve with greensand application; however, because of its slow release of potash, large quantities were required. Certain greensands that contain sulfur and sulfide minerals are harmful to plant growth, and these were classified as poison, burning, or black marls. The availability of higher grade potash salts from other mineral sources and the manufacture of prepared fertilizers displaced the agricultural use of greensand during the latter 1800s. During the mid-1800s the greensand industry, centered in a small section of the eastern United States, grossed more than $500,000/y. Toward the end of the century, however, annual production had dwindled to less than $100,000 in value. By 19 10 there were only six or eight greensand producers grossing less than $5,000/y each (Tyler, 1934). There was a brief revival of the US industry during World War I because of the curtailment of foreign potash, especially from Germany. During the latter 1940s and early 1950s greensand was again recommended as a food nutrient for plants and farm crops. Agronomic studies discussed its potential as a soil additive that gradually releases potash and many trace element nutrients essential for plant growth (Tedrow, 1957). Greensand was sold with the idea that it would condition soil and absorb and hold water while its base exchange properties would release trace elements. For a short time glauconite was used in certain parts of New Jersey as a binding additive in the brick industry, and in the 1800s it was used for making green glass (Cook, 1868). In the early 1900s the base exchange properties of glauconite were recognized for water treatment and the mineral gained acceptance as a water softener. Mansfield (1922) does not mention base exchange even though this phenomenon was known in 1916 or earlier. From 1916 through 1922 several patents for the use of glauconite as a water softening agent were granted. A method was also patented for treating greensand to improve it for water softening and ready regeneration with common sodium chloride brine (Borrowman, 1920, Spencer, 1924, Kriegsheim and Vaughan, 1930). Treated glauconite, on contact with water containing magnesia or lime, takes up magnesium or calcium ions and releases sodium ions. This exchange is limited to the outer surface of glauconite grains, and when all the surfaces have absorbed their capacity, the grains must be regenerated. Regeneration, simply stated, consists of treating or backwashing the glauconite with a sodium chloride solution, which replaces the hard water elements with sodium, thus reviving the glauconite. The process has become more sophisticated due to competition among companies in the water softening business. Greensand products for water softening generally consisted of several different grades distinguished by the particular treatment the glauconite was given during processing. The standard greensand water softener was produced from natural glauconite that was only washed and classified. Its characteristics for water softening are given in [Table 1].

Jan 1, 1994

By George S. Austin

Mud is one of the oldest building materials used by man, but it is not only of historic interest. Even today in many parts of the world, particularly lightly forested areas, it is the chief building material. Many authorities believe that at least 30% and perhaps 50% of the world's present population lives in earthen dwellings (Dethier, 1985, Coffman et al., 1990). Not all of these people are members of primitive societies or live in third world countries. European, African, Asian, and both North and South American countries contain a large number of such structures. In the United States, the American West, and particularly the Southwest, is known for this type of construction. Some of the earliest remains of adobe structures are those discovered in the ruins of Neolithic farming villages in Mesopo¬tamia dating as far back as 7000 BC (Steen, 1972). The word adobe has its roots in Egyptian hieroglyphs denoting brick and evolved to its present form through Arabic and Spanish (Lumpkins, 1977). The Spanish conquest of the New World spread the use of wooden molds to produce a standard adobe brick. Today, the word adobe is used to describe various earth building materials and techniques, usually referring to sun-dried adobe brick now used in the United States, but is also applied to puddled adobe structures, mud-plastered logs or branches (Jacal or waddle-and-daub), pressed-earth blocks, and rammed-earth walls or pisé (Smith and Austin, 1989, Ferm, 1985). Modern mud construction is used in many countries in various parts of the world. In the United States, the states from Texas to California are perhaps best known for this type of construction. Of these states, New Mexico has the dominant reputation for adobe use. Indeed, in New Mexico the Santa Fe style (Fig. 1) has made adobe not only acceptable, but preferred. RAW MATERIALS Adobe soil used by present day adobe producers, and probably past adobe producers as well, is principally sandy loam (50% clay and silt), although clayey silts are used in some areas (Coffman et al., 1990). In New Mexico, the best adobe soils are those developed on stream deposits, particularly Holocene (Recent) terrace deposits and older, loosely compacted geologic formations, such as the Santa Fe Group (Tertiary) located in the Rio Grande Valley. Some modern adobe producers use a mixture of materials from the screened fines of aggregate operations and mud from irrigation ditches in the river valleys combined with varying amounts of sand to produce the proper blend. Mineralogy Bulk Mineralogy: X-ray diffraction analyses of whole rock samples from many parts of the world where adobe is the dominant construction material show the major constituents of adobe are quartz and feldspar, with lesser amounts (in order of abundance) of calcite, clay minerals, and gypsum (Coffman et al., 1990). Adobes from and climates contain considerable calcite; in some cases calcite is second only to quartz in volume. Quartz, feldspar, most of the clay minerals, and some calcite commonly are derived from the mechanical/chemical breakdown of older rocks units. Some of the clay minerals, much of the calcite and perhaps all of the gypsum are precipitated from evaporating water. Clay Mineralogy: Clay-size minerals consist dominantly of clay minerals, but nearly clay-size fractions contain minor amounts of quartz and calcite, and occasionally other nonclay minerals. In earth construction materials from New Mexico, clay-size particles are the most compositionally variable in commercial adobe soils. However, the clay mineral groups in this size fraction consist of about equal parts of expandable clay minerals (smectite and mixed¬layer illite/smectite or I/S) and non-expandable clay minerals (ka¬olinite, illite, and chlorite), with minor quartz, calcite, and feldspar (Smith and Austin, 1989, Austin, 1990). In the and American Southwest, smectite is commonly calcium-rich and the I/S is dis¬organized and randomly interstratified. In a study of 42 New Mex¬ican commercial adobe soils, only two contained chlorite, and vermiculite, sepiolite, and palygorskite were not found. Soils in temperate climates, as are found in the midwestern and northeastern United States, contain an abundance of illite (Potter et al., 1975). The clay minerals in earthen structures of that region and in Europe commonly contain illite with less amounts of kaolinite, smectite, I/S, chlorite, and vermiculite. In contrast, the clay-size fractions of soils of humid tropical areas are typically acid, with kaolinite as the dominant clay mineral and lesser amounts of I/S, illite, smectite, and others (Chamley, 1989).

Jan 1, 1994

By L. H. Munson, D. E. Hadlock, L. F. Munson, R. L. Gilchrist, P. D. Robinson

All uranium mills are required to perform sampling and analysis for radioactive particulates in their gaseous effluent streams and in the environment. Pacific Northwest Laboratory was requested by the U.S. Nuclear Regulatory Commission (NRC) to provide technical assistance to them for their Uranium Mill Health Physics Appraisal Program. In conducting appraisals, air sampling methods used at NRC-licensed mills were reviewed and several deficiencies noted. This paper includes only environmental and effluent particulate sampling although much of the information is applicable to both in-plant and environmental samples. First, the components of a proper sampling program are discussed: program objectives, program design, sampler design, analyses, quality assurance, and data handling. Then the specific deficiencies, or the "pitfalls" from the first 8 mill appraisals are discussed. The first consideration in establishing an air sampling program is defining the objectives of the program. What is air sampling suppose to accomplish? Many of the deficiencies we have observed have resulted because the desired objectives were not clearly established in the minds of the radiation safety staff. PROGRAM OBJECTIVES An environmental air sampling program ought to fulfill the following seven objectives. The first is to: 1) [demonstrate regulatory compliance]. Although a goal of most programs, regulatory compliance, is not well understood. One has not only to comply with the conditions of the source materials licensee, but one must also demonstrate compliance with 10CFR20 and 40CFR190. For example, 10CFR20.106 states: "A licensee shall not possess, use, or transfer licensed material so as to release to an unrestricted area radioactive material in concentrations which exceed the limits specified in Appendix B, Table II of this part .... For purposes of this section, concentrations may be averaged over a period not greater than one year." Even if a mill's license does not require sampling at the site boundary of maximum concentration, a sample may be necessary to demonstrate compliance with 10CFR20. Most mill personnel are painfully familiar with 40CFRl90.10, which states: "Operations.... shall be conducted in such a manner as to provide reasonable assurance that: (a) The annual dose equivalent does not exceed 25 millirems to the whole body.... of any member of the public as the result of exposures to planned discharges of radioactive materials, radon and its daughters excepted... from uranium fuel cycle operations..." This means a licensee's sampling program must give "reasonable assurance" that the member of the general public receiving in the most exposure gets no more than 25 millirems per year. The sampling program necessary to provide that assurance may or may not be a license requirement. However, merely meeting the license requirements and the explicit regulatory requirements does not necessariarly ensure an adequate effluent and environmental air sampling program. The second objective of the environmental air sampling program, is to 2) [identify the source(s) of contaminants]. This will include not only the routine program, but special sampling for verification of sources and nonsources. Only after sampling can a mill operator be assured that roof vents, laboratory hoods, and other localized ventilation systems are not making a significant contribution to environmental releases. An environmental sampling program should also allow the mill operator to fulfill the third objective, to 3) [estimate exposures]. Even before 40CFR190, a sampling program should have provided the mill operator with the information necessary to determine the dose to the "fence post" person, or at least to determine if doses were well below the 10CFR20 limits previously allowed. The program should 4) [detect and measure unplanned releases]. If there is a fire, a scrubber failure, or if a drum of yellowcake breaks open, measured releases will almost always be lower than conservative estimates. Whether or not a system to provide sampling during accidents is needed is almost always a cost-benefit decision. In general, uranium operations do not sample just in case an accident may occur. Yet they may decide on continuous air sampling in lieu of intermittant sampling partially because of the potential for accidents. Another objective of air sampling is 5) [to provide information on the effectiveness of control systems]. This is always a concern with new or modified equipment and may dictate sampling frequency in other situations as well. For instance, if a small leak in a bag filter cannot be detected by other means, then more frequent stack sampling may be indicated. A routine effluent and environmental monitoring program should also fulfill the sixth objective,

Jan 1, 1981

By Francis S. Kendorski

INTRODUCTION Drift design problems in caving operations are a re¬sult of the geologic factors contributing to the overall success of the system, of the engineering factors dictated by economic and technical considerations, and of ore production practices. Combining these factors, a rational underground sup¬port system of rock reinforcement, light steel channel section or welded wire fabric, and shotcrete can be de¬signed based on rock fracturing, rock load, abutment loadings, ground movement, expected repair and desired flexibility. The design concept uses the effect achieved by restraining, reinforcing, and maintaining some of the intrinsic strength of the fractured rock mass composed of interlocked blocks of intact rock and rock fractures. Three different examples of drift support design in hypo¬thetical mines using the caving system are given. Caving is a system of underground mining where ore is extracted by means of gravity after the ore body is allowed to fail by removing support from underneath. The rock mass of the ore body fractures and flows ver¬tically downward to let gravity do as much work as pos¬sible. Caving differs from many other mining systems in that blasting is used only to initiate the rock mass failure by removing the rock supporting the ore but not to break the ore itself. The initial movement of the rock mass dur¬ing failure and the consequent crushing and grinding during the continued movement serve to reduce the ore to particles of a manageable size, with only limited sec¬ondary blasting necessary. The broken ore is extracted from the bottom of the failed rock mass through funnels of some sort pre-excavated in the rock. Ore extraction must continue or the swell of the broken rock will even¬tually fill the cavity and stop further rock mass failure and movement. The excellent general discussion on block caving in the SME Mining Engineering Handbook (Julin and Tobie, 1973) adequately covers the principles and application of this type of underground mining. Many rock mechanics aspects of block caving have been covered by others (McMahon and Kendrick, 1969; Swaisgood, et al., 1972; Mahtab and Dixon, 1975; King, 1946) and will not be reviewed further. Maintaining the stability of production drifts is one of the most troublesome problems plaguing the mine manager in a caving operation. Many factors contrib¬ute to drift support problems, and identifying the causes of instability and producing a reasonable support design are two steps toward achieving stability consistent with the mine plan. This chapter sets forth a technique for the design of support systems for production drifts in caving opera¬tions. The basic support system elements employed are rock reinforcement, welded wire fabric, and shotcrete. Recognized as contributing to the design are the factors of rock load, additional load from mining activity, rock fracture characteristics, repair expected, and flexibility. It must be emphasized that the drift must first be stabilized as for a tunnel, and the additional strengthen¬ing for mining-induced loads cannot contribute to the initial premining stabilization, or the reserve of strength is used up. MINE PLANNING The efficient mine planning engineer not only must satisfy the economic, human, and environmental aspects of his task but must also consider the mechanical con¬sequences of his plan. The problems created for the mine by placing parallel drifts too close, by crossing drifts on different levels with inadequate, if any, separa¬tion, and by installing connections and crossovers in the haulage plan without regard for the effective spans cre¬ated, are only a few of the problems a mine planning engineer can create for himself and the mine. The effect on immediately adjacent mine areas when an area is caved is important to drift design because the removal of vertical support from a rock mass causes the weight of that rock mass to be shifted elsewhere. The adjacent rock mass will carry this load and reach a new equilibrium with the applied stress. The advancing front of stress increase that results from caving (and many other mining systems) is generally called the abutment load and is the increase in stress over the gravity or tec¬tonic stress that already exists, as shown in Fig. 1. In general, the abutment load will be similar in nature to the stress change found around an opening in rock and will be taken as causing an increase in the vertical prin¬cipal stress, due to an approaching caving boundary, so that

Jan 1, 1982

By G. D. Naugle, L. C. Atkinson

Introduction Deep open-pit mines invariably affect the local and regional hydrologic systems. Pit dewatering, occurring during mining operations, puts an obvious hydrologic stress on these hydrologic systems. However, post-mining hydrologic impacts resulting from the pit refilling with groundwater following the cessation of mining activity can also be significant. The prediction of the rate at which the post-mining pit will fill with groundwater is a critical aspect of assessing the long-term hydrologic impacts. Numerical groundwater flow modeling provides a method for predicting the groundwater refilling rate of the pit. The rate at which pit "lakes" fill depends on several factors: •the rate and duration of pit dewatering; •the depth and size of the ultimate pit and •the pre-mining hydrologic regime. These factors can be incorporated into a detailed numerical groundwater flow model that can then be used to assess the effects of dewatering and post-mining recovery on the local and regional hydrologic systems. A sufficiently detailed, numerical groundwater model provides the oportunity to: •account for complex geology near the pit; •assess the impact of active pit dewatering and •predict the long-term impacts of post-mining groundwater flow into the pit. A detailed groundwater model incorporating these items has been developed and applied at an operating open-pit mine. Developed by Durbin and O'Brien (1987), the three-dimensional, finite-element, groundwater flow model was used to represent the hydrologic system of an approximately 253-km2 (98-sq mile) area surrounding the pit. Historical groundwater elevation data, stream flows and meteorologic, geologic and geophysical data were used to establish the dimensions and initial conditions for the model. Steady-state conditions, representing the pre-mining local and regional hydrologic systems, were simulated using the initial conditions incorporated into the groundwater model. The groundwater model was then utilized to simulate various dewatering programs, to predict the filling rate and the groundwater depth in the ultimate pit once mining activities are complete and to assess the long-term impacts on the regional groundwater flow system. Development of pit lake model Groundwater modeling efforts were completed in two phases. The first focused on pit dewatering activities, while the second phase concentrated on the post-mining effects on the hydrologic system. The final estimates of groundwater elevations calculated during the pit dewatering simulations were used in predicting the post-mining recovery of the hydrologic system. The groundwater model was also modified prior to the second phase to account for the volume of rock removed during mining activities. To account for the actual volume of rock mined, the geometry of the post-dewatering model grid was modified to approximate the final pit geometry. The depth and width of the ultimate pit were divided into eight idealized stages that represented significant changes in the bench geometries. These eight stages were then introduced sequentially into the model according to the predicted water elevations within the pit. In this way, changes in the volume and depth of water within the pit were accounted for through time. Once the ultimate pit geometry was accounted for in the model, it was necessary to assign new hydraulic characteristics to those parts of the model grid (elements) that represented excavated rock. The solution of the numerical model requires that finite hydraulic conductivity values be assigned to the portion of the groundwater model that represents excavated rock. Therefore, the calculated groundwater elevations differ, somewhat, between the edges and the center of the open pit. These model-calculated water elevations at the edge and in the middle of the open pit represent the elevation of water that would occur in the pit lake. To minimize the error in the estimated level of water within the pit lake, the hydraulic conductivity was increased to a value that would: •minimize the predicted difference between the groundwater elevations across the open pit and •produce a numerically stable solution. Specific storage is the hydrologic parameter that accounts for the water produced by compaction of the aquifer matrix. To predict the groundwater volume that would flow into the ultimate pit, this parameter was assigned a value equivalent to the compressibility of water. This value of specific storage reflects the post-mining groundwater storage occurring as an open body of water. Additionally, a specific yield of 1.0 was assigned to the pit elements to represent the 100% porosity of the open pit. In

Jan 1, 1994

By K. Fuenkajorn, S. Serata

Discussion by S. Serata and K. Fuenkajorn Background Results of the above study in the August 1991 issue of Mining Engineering offer valuable lessons in the solution of cutter-roof problems. The original study plan was initiated by the discussion authors to solve the problems using the "stress control method" of mining (Serata 1976, 1982; Serata, Carr and Martin, 1984; Serata and Gardner, 1986; Serata, Gardner and Preston, 1986; Serata, Gardnerand Shrinivasan, 1986; Serata and Kikuchi, 1986; Serata, Preston and Galagoda, 1987) However, the plan and the planner were changed to the arrangement reported in the paper. The change was considered reasonable at the time due to the mine engineers' uncertainties about the stress control method. Consequently, the basic principle of the study was shifted from the original stress control method to the method described in the paper, which will be called the "yield pillar method" for the purposes of this discussion. The paper convinces the reader that the yield pillar method fails to solve the cutter-roof problems. This doesn't mean that the stress control method also fails. Actually the contrary is true, as discussed below. Limitation of the yield pillar method The paper illustrates clearly how poorly the yield pillar method performs in solving the problem. The reason for this failure is the lack of the protective stress envelope needed to stabilize the cutter roof. Unfortunately, the protective envelope cannot be formed properly without utilizing the stress control method of mining. Changing the pillar size does not make much difference in the roof stability. Stress measurement The key issue is how to form the global stress envelope to make the gate entries safe for production. Therefore, measuring the stress condition of the ground around the mine opening is critically important to solving the cutter-roof problem, regardless of the method applied. With regard to the stress measurement, there is a serious question as to the reported stress state of [6 i = -51.7 MPa (-7499 psi), G2 = -44.5 MPa (-6458 psi) and 63 = -30.8 MPa (-4465 psi)]. It is mechanically impossible to have such a stress state at any location in the mine ground since the known initial vertical stress [o,,] is less than or equal to 800 psi. There may be a large stress state in the [61] direction, but that is possible only at the expense of the [63] value. Having the above stress tensors in the mine is simply impossible. The questionable, reported stress values could be attributed to the application of the overcoring method, which tends to produce erroneously large stress values in the extremely nonelastic mine ground. Stress control method The paper should be considered as a major contribution demonstrating the limitation of the yield pillar method. At the same time, the paper does not disprove the stress control method. However, in comparing the paper with stress control studies conducted in other similar failing grounds, the stress control method appears to be able to solve the problem more effectively. Therefore it is advisable that the mine not give up its efforts to solve the problem. [•] References Serata, S., 1976, "Stress control technique - An alternative to roof bolting?," Mining Engineering, May. Serata, S., 1982, "Stress control methods: Quantitative approach to stabilizing mine openings in weak ground," Proceedings, 1st International Conference on Stability in Underground Mining, Vancouver, BC, Aug. 16-18. Serata, S., Carr, F., and Martin, E., 1984, "Stress control method applied to stabilization of underground coal mine openings," Proceedings, 25th US Symposium on Rock Mechanics, Northwestern University, June, pp. 583-590. Serata, S., and Gardner, B.H., 1986, "Benefits of the stress control method," invited paper, American Mining Congress Coal Convention, Pittsburgh, PA, May 7. Serata, S., Gardner, B.H., and Shrinivasan, K., 1986, "Integrated instrumentation method of stress state, material property and deformation measurement for stress control method of mining," invited paper, 5th Conference on Ground Control in Mining, West Virginia University, Morgantown, WV, June 11-13. Serata, S., and Kikuchi, S., 1986, "A diametral deformation method for in situ stress and rock property measurement," International Journal of Mining and Geological Engineering, Vol. 4, pp. 15-38. Serata, S., Preston, M., and Galagoda, H.M., 1987, "Integration of finite element analysis and field instrumentation for application of the stress control method in underground coal mining," Proceedings, 28th US Symposium on Rock Mechanics, Tucson, AZ, pp. 265-272.

Jan 1, 1993

By Lowell T. Harmison

I appreciate very much the invitation to speak with you and the opportunity of bringing you messages from both the Secretary of the Department of Health and Human Services and the Assistant Secretary for Health/Acting Surgeon General of the U.S. Public Health Service. I would like to take this opportunity to congratulate you (the organizers of this Conference) on identifying the critical issues in the field and assembling such a broad array of experts to address them. I would like to present a brief view of the emerging framework for health that puts into perspective some of the aspirations of the Administration and to highlight several points with regard to prevention and occupational health. The goals are: 1. To improve the overall health status of our people. (This has been and will remain the National policy regarding health.); 2. To engage the Nation in the important effort of enhancing public health. (This is not reserved exclusively for the activity of the Federal Government or for State Governments. Public health has to be a cooperative effort that brings together all of the people engaged in the process of serving the people.); and 3. To pledge that health care will not be priced out of anyone's reach because of inflation. (It is clear that there are major tasks of bringing about economic recovery in our country. One aspect of this effort is to guard against the cost of health care not being allowed to rise beyond the reach of persons who need that care.) "How will these goals be achieved and what must change in the delivery of health and medical care in our society?" There are a number of real issues as well as perceptions that adversely affect the attainment of these goals: First, The cost of medical care is soaring and the public, industry unions and other elements of our society are becoming concerned. (They recognize the problem and are demanding a solution.); Second, There is a growing concern about the priorities that have been set. (For example, the evidence that preventive interventions are the most effective approach is overwhelming, yet medicine has not yet given that a high priority.); and Third, There is the perception that physicians do too much to too many people at too great a cost and that too much and too costly technologies are used. In view of the perceptions, we all must accept some changes and the challenges that needed changes will bring. A month before the new budget went to Congress, President Reagan went on nationwide television and told the American people that, "It is time to recognize that we have come to a turning point and we are threatened with an economic calamity of tremendous proportion and the [old business as usual treatment can't save us. Together we must chart a new course]." Now eight months down the road from this and a long Spring and Summer of discussion both within the Executive Branch and in the Congress, many plans and programs and concepts have emerged. The new course has been charted and the turning point has been made. Business as usual has been put aside and the Administration's leadership has been stretched and tested in putting forth a better approach with the reality that money is tight and that old habits of delivering care are difficult to change. The Congress has now given us a look at a new health budget that takes into account some of the harsh economic realities and that does make allowances for the persistence of familiar behavior. Against this background, it is now possible to begin addressing ways to provide health services to people at a price the Nation can afford to pay. There are without question difficult decisions involved but the Administration is committed to supporting and improving health care in America. It has been the President's contention that one of the principal causes of the inflationary spiral in the country was the steady and indefensible growth of the Federal budget. The problem stems from the fact that we have been living well, but beyond our means for nearly 30 years. Now we are discovering that there is a bottom to the barrel after all. It is possible for our society to run out of things like energy (oil), water or money. The health bills must be paid -- by Government, by insurance, by parents or by someone. Each year with a bigger shopping list and more money to spend the Federal Government went into the marketplace to buy. This action altered the

Jan 1, 1981

By Susan T. Bagley, Kenneth L. Rubow, David H. Carlson, Bruce K. Cantrell, Winthrop F. Watts

Permissible exposure limits (PELs) have been established for gaseous pollutants, carbon monoxide (CO), carbon dioxide (CO2), nitric oxide (NO), nitrogen dioxide (NO2), and some gas-phase hydrocarbons emitted in diesel exhaust. There is, as yet, no PEL recommended for diesel exhaust aerosol (DEA), nor is there a standard method for sampling this aerosol. The University of Minnesota and the U.S. Bureau of Mines have collaborated to develop a personal diesel exhaust aerosol sampler (PDEAS) which utilizes size-selective inertial impaction and gravimetric analysis. During the field tests of this sampler, numerous air quality measurements were made in underground coal mines that use diesel equipment. The mine mean DEA concentrations for the five mines surveyed, determined with the PDEAS in the haulageway, was 0.89 mg/m3 with a standard deviation of 0.44 mg/m3. DEA contributed 52 % of the respirable aerosol at this location. In three of the mines filter samples were collected for DEAassociated polynuclear aromatic hydrocarbons (PAHs) and biological activity determinations. Two of the mines were also monitored for the major gaseous constituents found in diesel exhaust. In general, the PAH and biological activity levels were similar for all three mines, and indicate that up to 25 % of the haulageway concentrations may be contributed by outby diesel vehicles. Measured concentrations of CO, C02, NO, NO2, and SO2, were well below regulated levels. INTRODUCTION Diesel exhaust contains pollutant gases, such as carbon monoxide, carbon dioxide, nitric oxide, nitrogen dioxide, and gas-phase hydrocarbons, as well as DEA. Much of the health-related concern focuses on DEA and associated organic compounds (Watts, 1992a). A wide variety of these PAHs have been identified and some are known carcinogens and/or mutagens. The U.S. Mine Safety and Health Administration (MSHA) has proposed new PELs for these and other contaminants (MSHA, 1989). MSHA has also published an advance notice of proposed rulemaking to establish a separate PEL for diesel particulate (MSHA, 1992). The U.S. Bureau of Mines has collaborated with the University of Minnesota to develop and field test a PDEAS. The PDEAS is a three stage sampler based on the MSA' personal respirable dust sampler. It utilizes a respirable cyclone preclassifier followed by a 0.8 µm cut point impactor and afterfilter operating at a flow rate of 2 L/min. Respirable aerosol greater than 0.8, µm in size is collected by the impactor while DEA, less than 0.8 µm in size, is collected by the afterfilter. Hence, gravimetric analysis of the afterfilter permits measurement of DEA concentrations. This development and laboratory evaluation of the PDEAS were described previously by Cantrell (1990) and Rubow (1990). During field tests of the sampler, numerous air quality measurements were made in continuous miner sections of five underground coal mines that use diesel haulage equipment. These air quality measurements included levels of selected PAH and biological activity associated with DEA collected in the intake and haulageway areas of three of the five underground mines, and CO, CO2, NO, and NO2 in two of the mines. The objectives of this paper are to present the DEA and associated pollutant concentrations measured in these mines and to assess the impact of diesel face-haulage equipment on underground mine air quality. MINE DESCRIPTIONS The mines used for the PDEAS evaluation were designated J, K, L, N, and 0. Mines K, N, and 0 are located in the Western United States, while mines J and L are located in the East. Each mine produces high volatile, bituminous coal with shift production levels varying from 500 to 2000 tons/section. Seam heights varied from 1.5 to 3.0 m. Mines K and N use continuous mining to develop longwall panels. The others are strictly room-and-pillar operations using continuous miners. The number and types of diesel-powered vehicles used at these mines were described by Watts (1992b). Mines J, K, N, and 0 use diesel power to assist in a wide range of activities in addition to coal haulage. These included road maintenance, personnel and materials transport, lubrication, and welding. Mine L used only three diesel-powered shuttle cars to haul coal. SAMPLING AND ANALYSIS METHODS Aerosol Measurements Aerosol samples were collected in the mine portal area, the clean air intake to the continuous miner section, the haulageway one crosscut inby from the feeder breaker and belt, in the return airway, and on selected personnel. The haulageway sampling site was located near the point where the diesel-powered shuttle cars turn around to dump their loads. Additional respirable and DEA samples were collected and have been reported by Haney (1990).

Jan 1, 1993