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By Feng Cai

"As a new method to support gate roadways, rock bolting has been widely used in the big coal enterprises for its superiority. The disturbed stress field of surrounding rock and anchor pre-stress field were divided into 3 areas in consideration of s1, s3. The surrounding rock failure form and characteristic were analyzed in combination with parameter f, the failure mechanism was explained using the slip line theory, furthermore, the rock bolting mechanism and its arrangement were studied. After roadway excavation, the tensile stress area appeared at the surface of the surrounding rock. The failure type changes from shallow tension-shear form into deep compression-shear form and the failure shape is an “incense burner”, surrounding by high risk failure area in butterfly shape. The maximum principle stress in roof appears discontinuous distribution because of layers separation. The failure type of surrounding rock is intensity-stress environment-weak plane control model. The internal reason of roadway macroscopic failure is the displacement between the surface plastic area and the deep elastic region along the tangential direction of slip line. In condition of rock bolt support, there is one-way tensile stress area in small range at the surface of surrounding rock, inside is the bidirectional compressive zone in uniform distribution form, stress environment is improved by the superposition of anchor pre-stress and disturbed stress field, anchor rock bolt arrangement and pre-tightening force distribution is decided according to the distribution of f parameter. Finally, supporting method and parameter reasonability was proved by the roadway deformation monitoring, damage depth drilling peep test in the 8101 transportation roadway in a mine.Introduction The stability of roadway support is very important to the safety and sustainable development of coal mine. In recent years, the focus of China's coal mining industry began to move to West China, the thick-extra thick coal seam is developed in the West. The latest technology and equipment have been put in to service, the occurrence condition of thick coal seam and the increasing of productivity lead to the cross section of the roadway increasing continually, hence, it is more difficult to control the surrounding rock. At present, the most common way to support the roadway is rock bolt (cable) support (Wang 2007); Scholars have been deeply studied on the mechanism of rock bolt. It is reported (Wang et al 2012, Gao 2007, Zhang and Yuan 2006) that the influence of rock bolt on physical and mechanical parameters of coal and rock mass in the anchorage zone, It is believed that the rock bolt can improve the peak strength of surrounding rock, especially the post peak strength, and the failure mode of rock mass under the rock bolt support is analyzed. It is reported (Wang et al 2014, Xiao et al 2013, Li et al 2007, Yan et al 2012), according to different geological conditions, the failure mechanism of surrounding rock and supporting technology of the whole coal roadway were studied. Focus on the failure mechanism of roadway surrounding rock, different forms of high strength support methods were proposed, which achieved good results. It is reported (Wang et al 2012a, Zhang et al 2011, Wang et al 2012b, Kang et al 2010, Zhang et al 2010, Kang et al 2007) that the supporting mechanism of pre-stressed anchor was analyzed from the point of the distribution of the maximum principal stress in the pre-stressed anchor field, and the influence of the rock bolt support parameters and surrounding rock characteristics on the supporting effects were studied. Based on the latest research results, this paper analyzed the failure mechanism of surrounding rock of coal roadway, proposed the high strength anchor net cable combined supporting technology based on the feature of stress zoning distribution of surrounding rock, to provide reference for the coal roadway rock bolt (cable) support technology."

Jan 1, 2017

By G. T. Lineberry, W. J. Wiehagen

A paradigm shift As market economists, mine managers are interested in "staying in business" and even "prospering." Bottom-line results are the "order of the day." Concern for bottom-line results can invoke questions such as how much was produced, at what cost, and was anyone injured? While these questions will always carry everyday emphasis at the mine, they are fundamentally lacking in that they often appear, especially to lower levels of management and to the production worker, as pre-eminent concerns. Driven by market economics, many organizations have decentralized, outsourced, downsized, restructured and formed new coalitions. Investments in new technology and a concentration on productivity have been an essential part of these changes, contributing to improvements in both safety and efficiency. A reason for (and, perhaps a result of) these improvements has been the development of a highly experienced and flexible work force. This situation presents an interesting opportunity to modify a few paradigms about work life in general, and perhaps a few paradigms about how we invest in the miner, in particular. Studying these issues now night offer some interesting opportunities for the Future, as the foundations for the next generation of mine workers are laid. With the average "experienced" U.S. miner five to fifteen years from retirement, perhaps a key determinant of future success in the world market might be how well we make the transition from today's work force to the work force of tomorrow. We can ill afford to wait a few more years to tap into the special knowledge residing within a veteran mining work Force. The Work Crew Performance Model (WCPM) is one approach to defining, capturing and transferring this expertise. The WCPM suggests a paradigm shift, moving from a focus on products (quantity) to a focus on the process (quality). It also involves a slight change in the process -- a different way to think and work. Of great importance, it entails a belief and a commitment to people within the organization; a commitment to collect simple (but interesting) data; exploration for insightful ways to plot and integrate that data; and finally the search for creative methods to reinvest that knowledge back into the organization, that is, back into the work force. As a way of thinking, the WCPM subordinates information about how much is produced daily or how many were injured over a year's time to concern and detail for how the individual's and the work crew's performance contribute toward meeting organizational goals. Drawing upon the expertise within members of that work crew to address and solve operational problems is one approach for meeting organizational goals. This everyday focus on how we work is hypothesized by the WCPM to make the difference between a good section supervisor and the average one; an exemplar continuous-miner operator and the run-of-the-mill one; a mining crew that produces consistently high quantities, but, inadvertently, degrades the production of other units. The WCPM permits objective and reliable performance data to be collected. This data then can be used to define variability within tasks of work crew members; to relate observed variability to a cost consequence; and to more meaningfully analyze performance through the integration of traditional analytical tools, such as the production simulator, CONSIM, with a recommended behavioral approach, the WCPM. Basis for engineering test Most equipment selections and purchases involve a field trial test of the system. Technology is routinely bench tested at the manufacturer's facility, government agencies and the mine site. The WCPM suggests a similar, but more practical, method for testing the performance of a work system -a test for how technology is used (by people) at the work site. It implies learning from veteran mining personnel that have gone well "beyond the book." This approach to learning and to integrating training with everyday work life can help answer questions such as: What constitutes desirable perfor¬mance as, for examples, a shuttle-car operator, a miner operator or a bulldozer operator; what characterizes exem¬plary performance within the context of the work system, the technology and the work crew; how can investments in the worker he linked to measurable results within the organiza¬tion? This paper explores these questions and makes recom¬mendations for enhancing the cost-linkages between invest¬ments in people and the mine's monthly cost sheets.

Jan 1, 1996

By Peter J. Huck, Yoginder P. Chugh, M. Jennings

INTRODUCTION Large areas of the United States coal reserves have been undermined by room-and-pillar mining over the past century. These abandoned mines generally cause subsidence of the ground surface many decades after mining. Reports from England (1) where there is a longer history of room-and-pillar mining indicate that mines abandoned two centuries ago may still be causing subsidence, and causing surface destruction to urban areas or farm lands that have been developed in the interim. The devastation wrought by subsidence, both in terms of property loss and societal disruption is enormous. The authors discuss here the methods by which subsidence in abandoned mines may be controlled. Two distinctly different subsidence control technologies have been developed in the U.S. for abandoned coal mines. These are referred to as (1) Point Sup- port methods, and (2) Areal Backfilling. As the names imply, point support methods are used to protect individual structures, or even individual foundation elements of structures, whereas areas backfilling techniques may be applied to protect entire neighborhoods or urban districts hundreds of acres in extent. Both point support methods and areal backfilling have specific sub-categories of techniques that may be applied under different conditions as appropriate. POINT SUPPORT METHODS Point support methods are characteristically employed by civil engineers chartered with the design of a new structure or the protection of an existing structure. The subsidence control measures are only a part of a larger project, the safety of which must be assured to a degree of near certainty. The costs of point support methods involve a large number of bore-holes and the use of expensive materials in relatively small quantities. There exist several dozen point support methods, which operate either to support the underground cavities, controlling subsidence, or to isolate the structure from the effects of subsidence if it should occur. The latter methods usually involve deep foundations through the mine opening supporting the structure on the underlying strata. Our concern will be with the former method to support underground cavities to prevent subsidence. Gravel Columns and Associated Methods Gravel columns may be placed in the mine opening using the methods shown in [Figure 1]. Boreholes are drilled through the mine strata, and gravel poured down the borehole to form a pile on the mine floor. When the tip of the pile contacts the mine roof, it is rodded down to spread the pile, and permit additional gravel to be placed. The objective is to place as much gravel as possible in the mine opening, and to achieve firm contact with the mine roof. The gravel pile acts to reduce subsidence by three mechanisms. The filling of a significant fraction of the open mine volume reduces the amount of potential subsidence simply by killing open volume within the mine. In effect, the extraction ratio has been reduced by introducing the volume of gravel. The toe of the gravel pile abuts against nearby pillars, providing lateral support and protection against pillar deterioration. In effect, the height of the pillars has been slightly reduced by burying their lower portions. Finally, some direct support is provided to the roof strata, reducing the roof span and enhancing stability. The gravel column method may be used in a variety of applications, and with materials other than gravel. For example, gravel columns may be placed in a close line around the perimeter of the building to form a continuous gravel wall. The interior of the site may then be completely filled with sand, slurry or other material. Alternatively, a lean concrete may be used in place of gravel, so that the piles achieve significant structural capacity. The grout column, shown in [Figure 2], begins with the installation of a conventional gravel column which is placed around a grout pipe that extends through the borehole into the mine opening. When the gravel column is in place, portland cement grout is injected in stages. At each stage, sufficient grout is injected to fill the interstices in the gravel to build a column perhaps two meters in diameter and a meter tall. The group pipe is then raised one meter, and after the previously injected grout has set, the next stage of grout is injected. The process builds a column of grouted gravel to the

Jan 1, 1982

By R. J. Lee

Asbestos analysis is conceptually simple. The objective, in most environmental analyses, is to determine the number of asbestos fibers per unit area or volume in a sample. The sample may be a consumer product, an ambient air sample, or a water sample. This determination is complicated by several factors: (1) sample collection and preparation, (2) counting techniques, and (3) operational definitions of asbestos in use. Three different definitions have been used in the study of environmental samples and are summarized in Table 1. The analytical procedure and detail required in the analysis vary with the definition of asbestos. Accordingly, this may be the only type of environmental analysis where the definition of a substance dictates the choice of analytical methods, and the rigor demanded of the analysis. Lead, arsenic, or beryllium are sub- stances having a constant definition and the method of analysis chosen depends on the concentration and precision required. However, the meaning of the term asbestos fiber is dependent on the eye of the beholder. Macroscopically, asbestos has certain physical properties that make it highly useful. Rockforming minerals do not display similar characteristics. Now, the question arises as to how three definitions of asbestos came into being. This is a result of two factors: (1) the definition of asbestos on any scale intrinsically involves the morphology or shape of the material, and (2) the advent of the transmission electron microscope permits detection of asbestos fibers and other mineral fragments in a size range where the physical properties of asbestos such as tensile strength, flexibility, or thermal resistance are difficult, if not impossible, to measure. There is a good deal of controversy about which definition is most appropriate for regulatory purposes because of a lack of appropriate health effects data (Cooper, 1978). Most differences of opinion center on the questions surrounding the following: 11 Morphology-Is a 3:l aspect ratio an appropriate medical or physical lower limit for the definition of an asbestos fiber? 2) Length-Are asbestos fibers shorter than some arbitrary length (5 pm) as hazardous as longer fibers? 3) Mineralogy-Is the mineralogical distinction between asbestos and nonasbestos minerals medically significant? While these questions are directly resolvable only with the biological and health effects data, ultimately such data reflects the ability of analytical procedures to fully characterize the materials used in health effects studies on animal models, and to monitor the exposure to particulates with different physical attributes in the workplace or community (Cooper, 1978). The following sections outline the analytical problems involved in distinguishing the asbestos and nonasbestos varieties of the same mineral; the analytical procedures available for identifying asbestos and measuring its concentration in environmental samples; the ability of these methods to distinguish asbestos fibers from nonasbestos particles; and their precision and reproducibility. Finally, we suggest some future directions in terms of definition and procedure that would greatly improve the quality of such analyses. ASBESTOS VS. NONASBESTOS Asbestos that is mined commercially belongs to two mineral families; amphibole and serpentine (Zoltai, 1978, Wylie, and Steel, 1978). The

Jan 1, 1986

By J. Le Gac, P. Zettwoog, J. Brenot, J. W. Piechowski, J. C. Nenot

INTRODUCTION The components of the radiological hazard in uranium mines are: 1. exposure to radon and its short-lived daughters 2. exposure to long-lived ore dust 3. external irradiation With usual ventilation rates, the dosimetry relative to the first component only requires to know each miner&apos;s exposure to a potential energy of short-lived radon daughters, (EXP[a]). Presently in France, (EXP [a]) is determined monthly from radon concentration measurements. This is an indirect method based on ambient monitoring. After technical improvements achieved during the last years it is now possible to measure (EXP [a] directly with an individual portable monitor developed by the CEA. The purpose of this study is to analyze and compare monitoring results obtained by these methods. DETERMINATION OF INDIVIDUAL MONTHLY EXPOSURES TO[ a] POTENTIAL ENERGY OF SHORT-LIVED RADON DAUGHTERS [(EXP a) amb, by ambient monitoring] In this case, it is necessary to know individual monthly exposures to radon, (EXP Rn). As mentioned by Pradel, J.,et al., 1976, (EXP Rn) is estimated for each worker from informations both on radon concentrations (CRn) measured as described by Pradel, J., 1962 and times (t) elapsed in the various workplaces. (EXP Rn) is obtained by summing up the elementary radon exposures CRn x t relative to all the workplaces where the miner was successively engaged during the month. (EXP Rn) values are then corrected by an average equilibrium factor of 0.17 deduced from many measurements and taken as representative for all the French mines. Finally: [ ] (EXP [a])ind are obtained directly with an individual portable instrument described by Chapuis, A.M., et al., 1972. Samples of the air breathed by the miners are pumped continuously and radioactive aerosols containing the short-lived radon daughters are trapped on a filter. The principle of the measurement lies on track-etch detection of the a particles emitted by the deposited material. The technical aspects have been described by Duport, P., et al., 1980a, Mouden, A., et al., 1980 and Chapuis, A.M., 1980. As mentioned by Duport, P., et al., 1980, fiability is good. Indications on measurement reproducibility are given by Pradel, J., et al., 1977. PRESENTATION OF THE STUDY The underground miner population of a French mine was followed during a period of 8 months (October 1978 - May 1979). Monthly exposure values for each miner determined either by ambient or individual monitoring, (EXP [a])amb or (EXP[ a])ind, had been previously collected by Le Gac, J., 1980, who had also defined miner categories according to the characteristics of the work performed. Four groups are considered here, the "whole mine" and three categories as described in table 1. Data on the individual monthly exposures,(EXP [a])ind and (EXP [a])amb, were taken again to analyze the individual cumulative exposures instead of the "month by month" values. The influence of the mean habits of each individual on its surveillance results is thus taken into account. Each miner has a cumulative exposure as an individual of the "whole mine" and as an individual of each category in which he has worked. The results obtained by ambient or individual monitoring were analyzed (characteristics and adjustements of the distributions) and then compared (description of the discrepancies observed). Obviously, such a study is possible only if all the exposures are cumulated on a same duration. As miners have not worked during the same number of months in the course of the data collecting period, one must considerer the individual average monthly exposures, [(EXP a)]amb and[ (EXP a)]ind, instead of the individual cumulative exposures, which is typically the same thing. To be sure that the exposure values of a given group (A, B, C or D) are representative, only those miners really present for at least 3 months in this group were taken into account. Finally, each individual of the "whole mine" and of the various work categories has a pair of average monthly exposure values : [(EXP a)amb] determined by ambient monitoring and[ (EXP a)ind] deduced from the track-etch recording of the individual portable monitor. Let a miner be engaged during n months[ (3 < n < 8)] in a group, the following relations then apply : [ ]

Jan 1, 1981

By C. J. Johnson, C. J. Bise

Coal extraction by continuous miners (CM) is currently the most common underground method in the US industry and accounts for slightly more than two-thirds of the nation&apos;s deep mining production (National Coal Association, 1987). Even if longwall mining should become more commonplace, it can proceed only after ventilation and access entries have been driven by CMs. Since an area of concern continues to be the effects of the dust generated on the health of mineworkers, this paper discusses the relationship between the elemental compositon of mining-generated airborne dust sampled from the immediate ventilation return of a CM and laboratory-generated dust derived from channel samples taken from the mines. There are several potential contributions of this type of study to the coal mining industry. First, after more fundamental knowledge of the causes of Coal Worker&apos;s Pneumoconiosis (CWP) is learned, the laboratory-generated respirable dust could be used to identify a potentially hazardous coal seam. Also, this study could possibly aid in understanding the fundamental causes of CWP by producing mining-simulated samples of coal dust that could be used in epidemiological studies. Further. assuming that there is no difference in the elemental composition of a drill-core sample and a channel sample from the same location, a mining company could predict anew mine&apos;s respirable dust elemental composition in the immediate ventilation return by using exploratory drill core samples of the roof, coal, and floor rock to prepare the laboratory dust. Ventilation engineers could then use engineering design and control measures during premine planning to reduce the incidence and severity of CWP by better ventilating the potentially hazardous coal seam. If this proper planning prevented any future changes to the ventilation equipment and mine design. much time and money could be saved. Scope of work To investigate the variability of the chemical characteris tics of respirable dust, airborne dust samples from eight underground coal mines located in the eastern and midwestern United States were collected with eight-stage Sierra Model 298 Marple cascade impactors, as well as 25 channel samples of mined material. Each channel sample was removed from the middle of the coal face before mining occurred. Sampling of the mining-generated dust was conducted by Lee (1986) by sampling the entire working sections, primarily for characterization purposes, to obtain information on the locational variability of dust characteristics. Research performed for this study used the elemental analyses of the mining-generated dusts he sampled in the immediate ventilation returns of CMs. The procedure that was used to produce the laboratory generated respirable dust was based on the Hardgrove grindability test since it reflects the pulverizing characteristics of coal. This test was chosen for several additional reasons. First, it is repeatable and reproducible. A consistent amount of input energy is used as well as a specified size range of feed material to be crushed (the channel samples). Second, it is thought to generate secondary dust in a way similar to that of the crushing and grinding of the coal and rock as they pass through the arc-shaped cutting path of the CM&apos;s cutter head. The potential effect on dust generation by this secondary grinding mechanism may be at least as much as that produced by primary fragmentation, which is dust produced by the cutting action of the bit against the coal or rock (Roepke. 1984). Finally, the Hardgrove grindability test is well known and is used in the coal industry to guide mineral processing engineers in estimating the capacity of mills used to grind coal. One hypothesis of dust researchers in the Generic Technology Center for Respirable Dust is that the elemental as well as the physical characteristics of coal mine dust will make a difference in the incidence and severity of CWP. Coal mine dust is generated not only from coal, but also from any rock partings contained within the seam or any roof or floor material mined with the coal. Thus, coal mine dust may not have the same elemental characteristics as the coal being mined. Given that hypothesis, mixtures proportional to each thick¬ness mined of roof, coal, and floor rock derived from the channel samples of the face areas from which the respirable dusts were generated by the CMs were used to produce the feed material that was pulverized in the Hardgrove machine.

Jan 1, 1990

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine&apos;s diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger&apos;s diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

By Scott Brien, O&apos

Cyclones have become standard unit operations in mineral processing for both dewatering mineral slurries and for making separations at given particle sizes. Conventional cyclones are normally able to make moderately sharp cuts. However, the limited ability to control water flow out of the apex means that cyclones have difficulty handling fluctuations in feed. Underflow-regulated cyclones are cyclones with a rubber boot attached to the underflow. This boot acts as a non-return valve and restricts flow out of the apex of a cyclone. Normally, this type of restriction on a cyclone causes severe difficulties in cyclone operation. However, by adding a siphon arm to the overflow, water flow can be controlled. The siphon arm creates a vacuum that reduces or eliminates the air core, permitting better control of the ratio of water flow out of the overflow and underflow. A Linatex-developed underflow-regulated cyclone was compared to a conventional cyclone for its ability to dewater a sand slurry. The underflow-regulated cyclone produced slurries with solids contents as high as 80% in the underflow. Conventional cyclones produced lower solids content in the underflow under similar conditions.

Jan 1, 1997

By R. Bruce Tippin

Background Iron ore is the No. 1 metal mining industry in the U.S. with dollar value of $2.3 billion in 1984 (U.S.B.M Mineral Commodity Sunnnaries , 1985). However, during the past decade this nation&apos;s iron ore industry has been subjected to a major market depression and a correspondingly downward adjustment in output. The recent trend in the curtailment of iron ore production traces a slow-down of the country&apos;s steel industry. Both pig iron and steel production have decreased significantly over the past several years. These trends are shown in Figure 1 from data collected by the federal Bureau of Mines (U.S.B.M. Mineral Commodity Summaries, 1985; U.S.B.M. Mineral Industry Surveys 1986). The industry is presently operating at less than 60% of its annual capacity. The domestic steel industry has been forced by reduced profits or losses to close facilities, curtail operations and restructure the financial status of several corporations. Companies have been sold or are trying to sell selected properties to improve their financial circumstances. Even with such actions, many of the steel companies are in very serious straits, including the seventh largest steel company, LTV, which has filed for bankruptcy. Many of the major steel companies have financial interests in iron ore mining and thus their adverse economic conditions directly reflect those operations. Several iron ore producers have been shut down including Reserve Mining Company in May, 1986 and Butler Taconite in June, 1985. The latter recently filed for bankruptcy under Chapter 11. A1 so in mid-1986, U.S. Steel Corporation, owner of the Minntac mine and iron ore processing plant, underwent corporate restructuring. The effect on their Minnesota plant is not known at this time. An excellent summary of the interrelationship of the iron ore companies and the steel producers has been provided by Skillings (1986), and an analysis of the iron ore situation was given by Robert F. Anderson, CEO of M. A. Hanna Company, in his keynote address at the 1986 University of Minnesota Mining Symposium (Anderson, 1986). Steel imports to the United States decreased slightly in 1985 because of import restrictions, but the long-term import situation remains dim and uncertain. As shown in Figure 2, the imports averaged about 25% in 1985, and the preliminary indications are that this figure could be as high as 30% when the final 1986 information is collected by the U.S. Bureau of Mines. At best, the industry can only hope for imports to stabilize at a constant level in the near future. Although the tonnage is small, the quantity of U.S. export steel has fallen over 50%. With many other materials replacing steel , the projected demand through 1990 is expected to increase only about 1% per year. Consequently, 1986 U.S. iron ore production will probably be 15% lower than in 1985. The 41 mil lion tons of iron ore production expected in 1986 represents only 53% of the industrial capacity, which is about 74.5 mil lion tons. Over 95% of this iron ore is in the form of beneficiated pellets. Today there is not an iron ore producer west of the Mississippi River, nor is there any production in the South. The Birmingham (Alabama) iron ore industry has been shut down since 1971. The western producers ceased operations in the early 1980&apos;s. Only the taconite operations in Minnesota and the plants in the Upper Peninsula of Michigan remain as our major domestic iron ore source. The economic situation for both the iron ore producers and the steel industry can be described as confused and in turmoil. Such a condition directly impacts the iron ore processing plants&apos; operations and plans for the future. Plant Practice At present the nation&apos;s eight major operating iron ore mines, listed below, are concentrated in northern Minnesota (Mesabi Range) and the Upper Peninsula of Michigan (Marquette Range). The only exception to the Minnesota/Michigan location is the Pea Ridge Iron Ore plant in Missouri, which is a subsidiary of St. Joe Mineral s.

Jan 1, 1986

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&apos;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 Friedrich Steinhäusler

RADON SOURCES AND CONTROL MEASURES IN THE MINING ENVIRONMENT Most of the contamination of the mine atmosphere by radon 222 is due to radon emanating from solid or fractured ore surfaces of walls, roof and floor. Also radon gas emanates from broken ore either from storage in backfilled mined-out areas as applied in e.g. shrinkage stopping methods or from ore spillage along intake airways mainly due to the use of trackless haulage. To a lesser extent water itself can represent an additional source of radon, which emanates into air from open drainage ditches or seepages along intake airways. The contribution from water can be controlled effectively by isolating the water from the primary intake air system, e.g. by diverting the water through pipes and/or sealing of seepages by grouting. However, control of radon emanating from rock surfaces creates a major technical problem with significant impact on the economic aspects of mining operations, if adequate radiological conditions must be maintained. Basically this can be achieved by suppressing the emanation process itself, confining already emanated radon or by removal of radon from the mine atmosphere. Extensive research has been carried out on the rate of radon emanation as a function of barometric pressure changes (Pohl-Rüling and Pohl, 1969). It could be shown that the radon supply consists of a permanent and variable component. The former results from the surface of the rock and depends mainly on the emanating fraction of its radium 226 content; the latter originates from within the rocks and is a function of the suction effect of decreasing barometric pressure, rock porosity and fissures. The practical application of this barometric pump effect for depressing the rate of radon emanation, e.g. by pressurizing the mine atmosphere, is limited due to high costs for providing a sink for absorption of radon and air as well as lack of permeability in most uranium ore bodies (Schroeder et al., 1966). Mine air cleaning by removal of radon can be achieved with the use of cryogenic methods, chemical removal, adsorption into charcoal beds, use of a gas centrifuge or general ventilation techniques. Technical problems have so far prevented the application of any of these methods other than ventilation. It is common practice to use the age-of-air concept, i.e. fresh air is delivered to the worker as directly as possible and removed quickly afterwards thereby maintaining the air "young". Engineering principles for quantity distribution of air through underground working areas are straightforward for general mining situations where radon constitutes an environmental contamination problem. However, in cases of high uranium ore content this concept may result in high costs with regard to installation and energy requirements for effecting both frequent air changes as well as sufficient heating of the air in cold seasons. Taking into account that the investment in ventilation systems is a major cofactor for the overall ore production costs this can be a limiting and decisive component in the assessment of the economic feasibility of specific mining operations and mineral reserves in general. Effective control of the radon flux from the rock surface prevents the initial contamination of the mine air with radon directly at the source. A radon diffusion barrier for practical application in mining requirements should fulfill the following requirements: - reduction of radon emanation rate by at least an order of magnitude - high mechanical strength - ease of sealant application onto surface to be coated - water resistant - low fire hazard - resistant to temperature changes encountered in mines - high cost efficiency in relation to exposure reduction achieved (direct and indirect costs) - low degree of maintenance. In the past several materials have been tested as sealants for controlling the emanation of radon from surfaces of rock and building materials. Epoxy paints reduce radon emanation rate only by a factor of 2 to 6 (Auxier et al., 1974; Eichholz et al., 1980; Keith Consulting Engineers, 1980). Although it is possible to prevent the escape of more than 99 % of the radon to the environment with gel seals over 80 mm thick (Bedrosian et al., 1974), practical applicability is very limited. Multilayer coatings of epoxy resins with various additives require meticulous preparation and flawless application of seamless four-layer coatings in four days to impede radon diffusion (Culot et al., 1976), otherwise results from this method have not been totally satisfactory (Leung, 1978). Aluminium foil laminated with polyethylene and paper on each side is under test as radon barrier but results are not available yet (Ericson, 1980). However, this method has the inherent disadvantage that possible malfunctioning electrical installations can cause fire or electrical shock through the sealant. Polyurethane foam coatings have been used on stoppings as very effective sealants. It does, however, represent a potential danger of spontaneous ignition and it is expensive (Rock, 1975). Thus, there is still need for a material which has similar properties as outlined above. In the following results are reported from investigations on the suitability of various materials as radon diffusion barriers.

Jan 1, 1981

By Fred Leighton

INTRODUCTION Dynamic (seismic or microseismic) methods of determining the stability of structures in rock are based on detecting and analyzing the characteristics of seismic energy that has originated from or traveled through the rock mass. This seismic energy can be in the form of naturally occurring rock noise energy resulting from structural adjustments within the rock or can be introduced into the structure by physical means, such as by blasting or impact. In either case, the seismic energy radiating through the rock mass can be detected using standard equipment and can be analyzed by established techniques to reveal a wide variety of information concerning the condition and stability of the rock mass through which the energy has traveled. In the following sections, the basic instrumentation required for seismic and microseismic studies is described, and some of the presently used applications of these methods are discussed to exemplify the state of the art. INSTRUMENTATION Seismic disturbances in a rock structure generate two types of seismic wave radiation, body waves and sometimes surface waves, which radiate outward in all direc¬tions from the source of the disturbance. Underground mining applications are generally concerned only with discerning the characteristics of the resulting body waves, i.e., the compressional (p-wave) and the shear (s-wave) energy. As these two forms of energy travel through the rock structure, the particles of the rock mass are caused to vibrate, and the vibration character¬istics resulting from each of the two types of wave are distinct. Some important differences are: 1) Compressional and shear waves travel at different velocities through the rock structure. 2) The frequency at which each wave causes particles to vibrate is different, and may range from about 50 to 100 000 Hz. 3) The amplitude or energy level of each wave is different, with the shear energy usually being the greatest. These differences form the basis for equipment se¬lection for individual studies and for modern data analysis techniques. The following sections describe the basic equipment necessary to detect and record seismic wave energy data and show several examples of analysis procedures and how these procedures have been used. In principle, seismic equipment is very simple. It consists of a geophone (or geophones) to detect the seismic energy vibration and convert that vibration to an electric signal, an amplification system to increase the level of that signal, and a means of monitoring and/or recording the signals detected. Fig. 1 is a block diagram of a typical system. The following sections offer a very brief discussion of system components and their individual functions. A more complete discussion is given by Blake, Leighton, and Duvall (1974). Geophones The function of the geophone is to detect the vibrations caused by the passing of the seismic wave energy and to convert that vibration into an electrical signal that displays both the amplitude and frequency characteristics of the vibration. Particle motion or vibration can be quantified and measured by measuring displacement, velocity, or acceleration of the particles. Thus, there are three types of geophones: displacement gages, velocity gages, and accelerometers. The choice of gage depends on the characteristic frequencies of the seismic energy to be monitored and the sensitivities of each type of geophone. In general, displacement gages are used for low-frequency monitoring (periods to 1.0 Hz), velocity gages for medium-frequency monitoring (1.0 to 250 Hz), and accelerometers for high-frequency monitoring (250 to 10 000+ Hz). Experience has shown that in underground studies, the choice of which gage to use lies between velocity gages and accelerometers. An easy, accurate method for selection of gage type is discussed by Blake, Leighton, and Duvall (1974). Once the type of geophone has been selected for use, it must be properly installed, and in the installation procedure the most important step is insuring that the gage is firmly attached to a competent portion of the rock structure. Poorly mounted geophones may entirely fail to recognize low-level seismic signals and will distort the information from signals they do see. Amplifiers Seismic events associated with mine structures occur over a very broad range of energy which results in a broad range of geophone output levels. In general, geophone output levels occur in the microvolt to low milli-volt range, and it is necessary to amplify these signals in order to drive recording or monitoring equipment. Because either an accelerometer or a velocity gage might be used as the geophone, the amplification system must

Jan 1, 1982

By W. E. L. Minter

The most important witwatersrand gold placers occur on extensive unconformity paleosurfaces that were produced during protracted periods of fluvial degradation. The erosional etches on what was generally unlithified substrate indicate that low sinuosity erosion patterns prevailed. The transport systems were dominantly braided fluvial with a minor debris-flow component. The placer sediments were non-cohesive sands and gravels which were evidently transported by fluctuating, yet perennial, discharges across extensive, low relief, bra1dplains, not far from littoral environments. Progress made in understanding heavy-mineral transport processes now emphasizes the importance that selective entrainment probably had in producing concentrations. Cognizance that these characteristics are not unique to the Witwatersrand should be taken into account when designing exploration models.

Jan 1, 1988

By A. T. Yu

Once exploration is complete, "winning" of the ore can be viewed as fundamentally a bulk handling proposition. Whether it is to strip, break, size, extract, load, haul, or even process, the success of each of these phases will depend upon how effectively one can utilize-the methods and tools that are available in bulk handling. Since as much as 30 to 60% of the total delivered price of raw materials can be attributed to handling and shipping (ref. 1), this expenditure may spell the success or failure of an entire mining venture. In fact, it has become increasingly evident that handling and shipping are so intimately associated with mining that only when they are duly considered throughout the mine-to -smelter master plan as an integral part of a total system can an optimum operation be expected and the most economical ton of ore obtained, this total systems approach should not only take in handling and haulage in and out of the mine, but also storage and transfer at the terminals (ref. 2,3,t), including ship and barge loading and unloading which are likewise a major consideration (ref. 5). Much has been developed during the past century in bulk handling, particularly in its expanding use in the mining industry. In view of the major role bulk handling has played and will continue to play in mining, in commemorating AIME&apos;s centennial, this paper will attempt to outline the state-of-the-art at this three-quarter point of the 20th century - one that is marked by its industrial revolution and rapidly accelerating technological advances. Bulk handling systems Although many developments in bulk handling are worth noting. The most significant during the past three-quarters of a century are: the total systems approach to solving problems, automation, the increase in speed and capacity.

Jan 1, 1971

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

1.1 GENERALIZED METHODS OF SUPPRESSING THE INFLOW OF GROUND WATER DURING THE CONSTRUCTION OF SHAFTS, DRIFTS AND TUNNELS Various methods are used to prevent or minimize the inflow of ground water into underground workings during their excavation. The two most common methods include freezing the saturated rock and grouting using cement, sodium silicate, polyurethane and/or other chemicals. Each of these technologies for combatting the inflow of ground water is effective only under specific hydrogeological conditions. For example, although the freezing of saturated ground is among the more universally adopted methods, it is designed to provide only temporary protection during construction. Before the saturated rock thaws it is necessary to emplace a waterproof liner which is labor intensive, time consuming, and expensive. Consequently, freezing is used only in exceptionally complex hydrogeological conditions, namely in those cases where the water-bearing strata consist of unstable rock or the ground water has an anomalous hydrochemistry. The classical grouting of saturated rock, carried out from the surface or from the face of the underground workings for the purpose of limiting the inflow of ground water during excavation, utilizes both cement and a variety of chemical grouts. Cement grouting has been regarded as the main method of combatting the inflow of ground water in fractured rock throughout the world. In such countries as the Federal Republic of Germany (FRG), Canada, the Republic of South Africa and Great Britain, cement grouting is the main method of reducing the inflow of ground water during the excavation of mine openings. In Great Britain and the FRG, cement grouting has been used in 80% of all shaft excavations. In the Republic of South Africa cement grouting has been used in almost 100% of all the shafts that have been constructed. The installation of grout curtains into permeable water- bearing strata significantly reduces their permeability and increases the rock strength. Grouting has the greatest effect in fractured sandstones, certain well indurated shales, fractured granites, fractured quartzites, and karstic limestones or dolomites. The following principal factors must be considered when assessing the expediency of grouting rock with cement: the geometry of the network of fracture openings, the saturated hydraulic conductivity of the fractured rock, the hydraulic head acting on the water in the rock, and the chemical composition of the ground water. On the basis of the geometry of the fractures and the thickness of each hydrostratigraphic unit, the characteristics of the cement grout are selected. As a rule, the cementing of large fractured zones with high ground water velocities is carried out using inert fillers (sand, mill slag, loam, loess, crushed limestone), special types of cement, setting accelerators, and high cement concentrations in water. Calcium chloride, soda ash, sodium silicate, sodium nitrate, amino alcohols, tin bichloride, trisulphate nitrate, and lumnite are used extensively as the setting accelerators for cement grouts. In the FRG and Poland, special cement injection compounds that embody a mixture of cement and active cement metal salts, water additives and binding substances are being used for the cementing of saturated zones with rapid ground water velocities. These reagents accelerate the rate of structure-forming reactions. In spite of the wide variety of additives used for cement grouts, the effectiveness of the method in large fractures below the water table is either poor or it leads to a very large consumption of cement due to the erosion of the grout through the cracks before it hardens. At the present time, a large number of cement types and brands are being produced by various countries. These variations permit cementing to be employed in a variety of geological conditions. However, both pure cement grout and grouts with fillers constitute unstable systems with a high water loss rate. Therefore during the grouting of finely fractured rock the cement grout&apos;s premature loss of a large amount of free water causes its consistency to increase, whereupon the grout solidifies. Consequently, it inadequately penetrates into the fine fractures of the stratum. In addition the high water loss rate in the cement grout causes unreacted cement particles to remain, which greatly reduces the grout hardness and the resulting rock strength characteristics are weakened. As a result, the binding properties of the cement are utilized perhaps up to 60%. The remaining portion is left in the grout as a filler. All these conditions significantly decrease the efficiency of the isolation effort both with respect to strength and cost. In order to expand the cement grouting method, re- search is being conducted to improve the quality of cement grout, to increase its range of application with respect to permeability control, to reduce its water loss rate and to increase its capacity to withstand the erosive and corrosive effects of poor quality ground water. For example, mixtures of bentonitic clay, carboxymethyl cellulose, aerated sodium sulfide, sulphated alcohol distillery waste, nitrolignin, gip-

Jan 1, 1993

By V. A. Leach, Lokan. K. H., S. B. Solomon, R. S. O’Brien, L. J. Martin, K. N. Wise

INTRODUCTION The development during 1979 of a relatively small, but high grade (10,000 tonnes uranium at an average grade of 2 per cent), uranium ore body at Nabarlek in the Northern Territory, Australia offered an excellent opportunity to obtain detailed radiation data for an open cut mine operating during the dry season. The ore body (Queensland Mines Limited-1979), which was completely extracted in a period of four and a half months, consisted of a vein type deposit dipping at 30 to 45 degrees and contained a central core of pitchblende in massive and irregular pods, surrounded by lower grade fine grained disseminated pitchblende. Mineralisation extended from the surface to a depth of 72 metres over a length of 230 metres with an average but variable thickness of 1D metres. Ore near the surface had been heavily weathered and complex secondary minerals were formed which had dispersed from the main vein. Mining was carried out with large earth moving equipment. Overburden and weathered surface ore were removed initially with scrapers. At greater depths bulldozers were used to rip and assemble ore and rock at each level, and these were removed by large trucks to the ore and waste rock stockpiles. Where necessary, blasting took place during shift changes each evening. Mining was essentially continuous with two ten hour alternating shifts working for thirteen days out of fourteen. At the completion of mining a relatively small excavation (335m x 185m x 70m) remained, and this will serve as a tailings repository during the milling phase. FIELD MEASUREMENTS The inhalation of radon daughters, arising from the radioactive decay of radon gas is well established (Archer et. al. 1973) as a potential hazard in the uranium mining industry. Control over radon and its daughters to ensure that recommended exposure limits are not exceeded is achieved by providing adequate ventilation, and under normal circumstances natural ventilation from an open pit should be sufficient. However, during the dry season it is not uncommon for stable atmospheric conditions, with little horizontal air movement, to develop - particularly at night - and significant radon daughter concentrations may accumulate. Throughout the entire mining period measurements were therefore made of radon and radon daughter levels at representative locations within the pit and on the ore stockpile as it developed. Initially these measurements were carried out manually, using the Rolle method for radon daughters, (Rolle 1972) and .scintillation cells (Lucas 1964) or a two filter tube for the determination of radon (Thomas 1970). For the latter half of the period however, a continuous recording instrument, developed within the Laboratory was used to provide a detailed record of radon daughter levels within the pit. At the same time, continuous readings of wind speed and direction, and vertical temperature gradient between 10 and 3D metres were recorded on a 30 metre meteorological tower, situated 800 metres from the pit. Radon Emanation Rates It is evident that radon and radon daughter concentrations depend on the grade, or more particularly, on the surface radon emanation rate of the ore which is exposed. Accordingly, as the mine progressed, detailed measurements were made of both of these quantities. The surface emanation rate of radon was determined for each ore bench as it was exposed by placing an extended array of canisters, filled with freshly degassed activated charcoal, face down on the ore for a known time. These canisters, which had previously been calibrated in the Laboratory, adsorb radon with high efficiency, and the total radon adsorbed is measured after retrieval by detecting the gamma rays from the trapped radon daughters (Countess 1977). At the same time, as each canister was placed in position, a measurement of the local ore grade was made for each location. This was achieved with a calibrated sodium iodide scintillation detector, adjusted to detect the 609 keV gamma ray from the isotope 2148i, a decay product of radium. Finally, measurements were made of the radiation field 1 metre above the surface, with a gamma ray survey meter, which was calibrated in the Laboratory. The relationship between the scintillator count rate and ore grade was determined by comparing the scintillator output with the gamma monitor, and relating the latter measurements to ore grade (Thomson and Wilson 1980). It was observed that while emanation rates and ore grades varied widely, the ratio of emanation rate to ore grade was in general fairly stable. A plot of this ratio is presented as a function of depth below the original surface in Figure 1. For most observations the ratio is constant at a value of 80 Bq m-2 s -1 per unit ore grade, where ore grade is expressed as percentage of U308. At the surface however, where the ore was weathered, the ratio was about a factor of three higher, and at two particular depths, where high grade pitchblende was being removed, it was very much lower. This was not unexpected as earlier Laboratory studies of drill core samples from Nabarlek had indicated that the emanation coefficient (the fraction of radon produced within the ore which escapes from the mineral particles) decreases with increasing ore grade.

Jan 1, 1981

By M. E. Breitrick

Newmont Gold Company began full production in early 1995 at the Carlin East underground mine. The mine is currently producing at the rate of 900 tpd (1.000 stpd). for an annual production of 150.000 ounces. Approximately one hundred twenty five hourly and salary personnel are employed at the project for the estimated six-year mine life. Two mining methods are currently utilized at the mine. Longhole stoping with cemented backfill accounts for 70% of the total mine production. Conventional underhand drift and cemented fill accounts for the balance of the daily tonnage. In December 1996. Newmont purchased a new AC-Eickhoff ET-210 Roadheader. The machine went into service in mid-June of 1997 and is used in both longhole stope topcut development and cut and fill production at a rate of 225 tpd (250 stpd). The Carlin East mine is operated by Newmont Gold Company. The mine is one of four underground gold mines in production in northeastern Nevada owned by Newmont Gold.

Jan 1, 1998

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&apos;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 Ian M. Loomis, Keith G. Wallace

The Waste Isolation Pilot Plant (WIPP) is a U.S. Department of Energy research and development facility designed to demonstrate the permanent, safe disposal of U.S. defense-generated transuranic waste. The waste storage horizon is 655 m (2150 ft) below surface in bedded salt. To date the WIPP project has not emplaced any waste. There are three intake shafts used to supply air to the underground. All air is exhausted though a single return shaft. The total design airflow during normal operations is 200 m3/s (424,000 cfm). The ventilation system is designed to provide separate air splits to construction, experimental, and storage activities. Separation is achieved by isolating the storage circuit from the construction or experimental circuits with bulkheads. Any air leakage must be towards the storage area of the facility. Field studies have shown that the pressure differential necessary to maintain the correct leakage direction is susceptible to the effects of natural ventilation; therefore, extensive studies and analyses have been conducted to quantity the natural ventilation effects on the WIPP underground airflow system. A component of this work is a monitoring system designed to measure the air properties necessary for calculation of the natural ventilation pressure (NVP). This monitoring system consists of measuring dry bulb temperature, relative humidity, and barometric pressure at strategic locations on surface and underground. The psychrometric parameters of the air are measured every fifteen minutes. From these data, trends can be determined showing the impact of NVP on the ventilation system during diurnal variations in surface climate. Both summer and winter conditions have been studied. To the author&apos;s knowledge this is the first reported instance of automatic and continuous production of time and temperature variant NVPs. This paper describes the results of the initial monitoring study. INTRODUCTION The ventilation system at the Waste Isolation Pilot Plant (WIPP) in Carlsbad, New Mexico, is designed to perform two distinct functions. First, it supports normal mine ventilation requirements complying with all state and federal mine regulations. Second, the system is designed to prevent an uncontrolled release of radioactive contaminants from the storage and transportation areas of the facility. Although a nuclear radiation release in the facility is considered unlikely, many special features are implemented in the ventilation system to prevent the possible spread of contamination. The facility is constructed with the waste transportation and storage areas separated from the mining and non-radioactive experimental areas. The ventilation system is designed such that air leakage is from the mining and experimental areas to the storage areas. Furthermore, radiation detectors are located throughout the storage and waste transportation areas underground and an exhaust filtration building is installed on surface to prevent the possible release of radiation to the environment. For over two years the underground ventilation system has been rigorously tested and balanced. It was during this period that the adverse effects of NVP were noticed and subsequently quantified. From extensive field studies and computer models, several mitigating features were designed and constructed and special operational procedures were implemented to control the impacts of NVP. To quantify more accurately the NVP at the WIPP, a continuous monitoring system was installed. This monitoring system consists of measuring dry bulb temperature, relative humidity, and barometric pressure every fifteen minutes at strategic locations on surface and underground. From this psychrometric data, the NVP is calculated. Fan operating pressures and flows and strategic differential pressures are recorded from the site Continuous Monitoring System (CMS). The monitoring system provides a means of evaluating how the ventilation system behaves in regard to climatic conditions and to judge the efficacy of the mitigating features and operational procedures. To the author&apos;s knowledge, continuous calculation of NVP as a function of time and surface temperature has not been previously reported. Overview of the Waste Isolation Pilot Plant The U.S. Department of Energy determined that the plastic nature of bedded salt may provide the best solution to isolate transuranic (TRU) waste from the biosphere. Initial evaluations at the WIPP site began in 1974. In 1979, the United States Congress enacted Public Law 96-164 for the construction and development of the WIPP project. The mission of the WIPP is to demonstrate the safe, long-term disposal of TRU waste generated by the national defense programs of the United States. TRU waste is classified as a low to medium level waste. The waste is stored in drums and does not produce significant heat (not greater than 1 W per drum). The WIPP site is located approximately 47 km (29 miles) east of Carlsbad, New Mexico in the Chihuahuan Desert. The repository is located in the 630 m (2000 ft) thick Salado Formation. This Permian Basin salt deposit is about 225 million years old and appears to have been minimally disturbed by earthquake, faulting, and ground water activity since it was deposited. The underground facility is 660 m

Jan 1, 1993

By N. S. Tanious

The Nuclear Regulatory Commission (NRC) regulation 10 CFR 60 requires that the ventilation system for the geologic repository be designed to: (1) control the transport of radioactive particulates and gases, (2) assure continued function during normal operations and under accident conditions, and (3) separate the ventilation of excavation and emplacement areas. The ventilation system design should take into account thermal loading; maintenance of the retrieval option and-the separation of the development and emplacement ventilation circuits. An analysis was performed by NRC&apos;s contractor, Engineers International on the ventilation system design described in the Department of Energy&apos;s conceptual design for Nevada. The ventilation analysis is of a preliminary nature. The conclusions presented in this paper reflect the views of the authors and may not represent the views of the NRC. In this analysis, available Nevada Nuclear Waste Storage Investigation (NNWSI) ventilation design information was reviewed and a ventilation network analysis was performed.

Jan 1, 1988