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By S. Tavani, R. E. Browner, R. Dunne

Lead is used in CIP/CIL circuit to increase the leaching rate of gold from ores. At least four mechanisms have been proposed for the action of lead in gold leaching circuits: (I) precipitation of sulphides as highly insoluble lead sulphides preventing consumption of cyanide and oxygen; (2) su,face reaction on sulphide particles forming an insoluble lead sulphide layer peventing dissolution of those sulphides; ( 3) precipitation of lead sulphide, which undergoes further oxidation to soluble lead sulphate recycling the lead to solution,· ( 4) contact reduction of lead onto the gold szaface which then acts as a secondary oxidising agent. Rotating disc studies using a pure gold disc show that small amounts (5ppm) of sulphide ions are sufficient to significantly reduce the leaching rate of gold. Lead increased the leaching rate at small concentrations but at higher levels decreased the rate of leaching. The detrimental effects of lead were only found for a pure solution when leaching the gold disc. Ores mediate the amount of lead available by ads01ption/reaction, thus, reduced leaching rates from excess lead are not serious even when 500ppm of lead ions are added to the sluny. The detrimental effects of sulphide ions could be countered by lead addition during the leaching process. A penalty, in the form of reduced leaching rate, will occur the longer lead addition is delayed. A time lag was noted between the addition of lead and the increased leaching rate when a large concentration of sulphide ions was present.

Jan 1, 2000

By Debra C. Man, S. Thomas Freeman, F. Wiley Horne

INTRODUCTION The Metropolitan Water District of Southern California (Metropolitan) is the largest water wholesaler in southern California. Its service area spans the major urbanized portions of Ventura, Los Angeles, Orange, San Bernardino, Riverside and San Diego counties. More than 15 million people are serviced by Metropolitan through its 27 member agencies. The primary sources of Metropolitan's water supply are the Colorado River Aqueduct (CRA) and California's State Water Project (SWP). Currently, planning and design level studies are underway by Metropolitan to expand the water storage and conveyance systems within the service area for the purposes of providing increased efficiency and flexibility to manage limited water supplies, improving water quality, and increasing system reliability. The system reliability issue is particularly significant because the conveyance facilities from both the CRA and the SWP must cross the active San Andreas fault. This fault is considered to be the source fault for the anticipated "Big Earthquake" in southern California. To meet these objectives, Metropolitan's expansion plans involve a major new storage reservoir, new large diameter pipeline and tunnel conveyance facilities, more efficient utilization of existing groundwater basins, and an aggressive public awareness and conservation program. This paper focuses on the planning studies that have been carried out for over 50 miles of tunnels being considered for several of Metropolitan's proposed conveyance systems. These tunnels have ranged in length from a few hundred feet to over 42,000 feet, with depths of cover ranging from a few tens of feet to over 3,000 feet, and finished diameters ranging from 10 to 20 feet.

Jan 1, 1993

By Ronald P. Geitgey

Pumice and volcanic cinder are volcanic rocks characterized by a cellular structure. They are formed as gases (primarily water) dissolved in molten rock, or magma, exsolve to generate a froth that cools and solidifies into rigid foam. The cells or bubbles are referred to as vesicles and they range in size from a few thousandth of a millimeter to several centimeters. Due to their vesicular character, both pumice and cinder have lower density and higher porosity than most other rock types, and, as the vesicle walls are broken, sharp cutting edges are continually generated. These properties are the basis for their commercial value as lightweight aggregates, insulators, absorbents, and abrasives. There is a certain imprecision in the terminology applied to volcanic products, reflecting to varying degrees the diversity of volcanic rocks, the relative youth of volcanology as a distinct field of study, and customs and practices of commercial trade. In general terms, pumice is light colored (white, gray, pink, pale yellow, or brown) and highly vesicular, usually with vesicle walls that are visibly glassy. Many pumice fragments are light enough to float on water. Pumice is readily cut by steel tools and some can be crushed by hand. Individual pumice fragments may range from 1 or 2 m in one dimension to c1 mm. Commercially, finer fragments are called pumicite or volcanic ash. In volcanology the term ash refers to any fragment c2 mm in size regardless of its composition. Cinder, or scoria, is dark colored (black, red, brown) with thick vesicle walls that appear dull or stony and that may exhibit iridescence. Cinder typically is heavier than pumice and has a higher crushing strength. In volcanology the term scoria refers to dark vesicular material throughout a wide size range while it is common commercial practice to use scoria for larger fragments (usually greater than 2.5 cm) and cinder for smaller fragments. GEOLOGY Composition and Physical Characteristics Most pumice is silicic ranging from 60 to 70% SiO2, that is, dacitic to rhyolitic in composition. Pumice of less silicic composition, including basaltic, occurs but it is less common and does not have extensive commercial use. Scoria and cinder are typically basaltic to andesitic with approximately 50 to 60% SiO2. Pumice and cinder may contain phenocrysts of feldspars and various ferromagnesian minerals that crystallized in the magma prior to eruption. Fragments of rock through which the magma has passed may be entrained in the melt and wall rock may be fragmented and admixed during an explosive eruption. Some scoria particles have lithic fragment cores. Phenocrysts and lithic fragments are relatively unimportant in cinder end uses but their presence may be detrimental in certain pumice products. Pumice typically has vesicles <1 mm in size separated by thin walls. Vesicle shapes include irregular, spherical, elliptical, and elongate to the point of being tubular with a silky appearance (Fig. 1). In some pumices, the vesicles are interconnected making the fragments permeable and highly absorbent; in others the vesicles are isolated forming a highly porous but very impermeable pumice. Floating masses of pumice fragments from the eruption of Krakatoa in 1883 were reported in the Indian Ocean for up to two years after the eruption. Vesicles in cinder are larger and range from spherical to highly irregular in shape with much thicker walls and abundant interconnections. Scoria and cinder are typically heavier, porous, and do not float. Density and hardness of pumice must be expressed precisely to avoid confusion. Density may refer to the glass itself, the apparent density of the vesicular pumice particle, or the bulk density of pumice in a deposit or in a product. Pumice glass has a specific gravity of 2.5 or more depending on its composition. Pumice fragments typically have a specific gravity of 4.0 and so will float on water, at least until enough vesicles are saturated to cause them to sink. Expressed as density, pumice fragments typically weigh less than 1 g/cm3. Depending on moisture content, particle density, and particle size distribution, the bulk density of pumice typically ranges from 500 to 700 kg/m3. Typical bulk densities of cinder range from 700 to 900 kg/m3 for material used as lightweight aggregate while the bulk density of cinder used for highways and railroad ballast may be considerably higher. Hardness in pumice may refer to the glass vesicle walls, possibly including crystals or lithic fragments, or to the apparent hardness of the particle as a whole, which is more a measure of the strength of the vesicle structure. Pumice glass typically has a Mohs hardness of 5 to 5.5. A pumice particle may have a much lower apparent hardness and be easily cut with a knife or steel saw because the vesicle walls break readily. Pumice and cinder are primarily pyroclastic deposits formed as fragmental products of volcanic eruptions. Several classification - systems differing in various details have been proposed for pyroclasts and pyroclastic deposits based on particle size (Fisher, 1961, Schmid, 1981, Cas and Wright, 1988). Generally, particles >64 mm are referred to as blocks or bombs depending on their shape. Particles between 64 mm and 2 mm (4 mm in some schemes) are called lapilli and smaller particles are called ash. These names may be modified by compositional terms, for example pumice lapilli or scoria lapilli. Most commercial deposits are composed of fragments in the lapilli size range. In the United States, pumice fragments having one dimension of 5 cm or more are legally defined as block pumice. The procedures for acquiring pumice from federal land are determined in part by this definition and are described in a later section.

Jan 1, 1994

By Rolf Pollak, Jan Tegtrneier, Gerhard Keller, Horst Gerhardt, Wolfgang Müller, Dieter Tetzner

INTRODUCTION Under certain geological and mining conditions mining near the surface results in an intensive destruction of the overlying rock strata. If the system of mine openings remains filled with air after having completed all mining activities because the flooding of the mines depending on the morphology of the mining area does not exceed a certain level, this circumstance might result in dangers for the public. This is particularly true where uncontrollable air- ways occur within the destroyed rock or if former uranium mining areas are concerned. In these areas radon emanation resulting in undue radon concentration should be taken into account. This circumstance can be counteracted by suitable measures of ventilation. In the following the example of a uranium deposit in the Erzgebirge Mountains (Saxon Ore Mountains) is used to illustrate how statements about the ground flow through overlying rock can be made on the basis of systematic investigations. Finally comparisons with experience gained in coal mining yield orientation data which are used as input quantities in ventilation network calculations and in project preparation of mine fans. Last not least it is proved that endangering the population by radon can be excluded by installing an appropriate fan. INITIAL SITUATION From 1946 to 1989 the uranium deposit of Schlemal-Alberode was mined starting from superficial zones down to depths of 1,800 m. The total output amounted to 80,000 t of uranium, i.e. Schlemal Alberode is one of the largest uranium deposits in the world. The mine openings comprise a cavity of 40 million m3. Al- though thin seams (0.2 to 0.3 m) and steep to semi-steep veins (60 - 70 degrees on average) were extracted using the method of overhand cut-and-fill there were also deposit sections with larger thicknesses and poor ore zones, respectively, where especially after the war until about 1955 large openings were formed due to missing filling. Due to this uncontrolled, relatively "wild" mining considerable subsidence of the surface and fractures in the border zone of the subsidence area occurred. The town of Oberschlema was destroyed and an area of 0.14 km2 was damaged by mining involving a maximum subsidence of 6 m in the centre - called deformation area. The rehabilitation work for the deposit area of Schlemal Alberode involves the flooding of the openings up to a level of the Markus-Semmler-Sohle as long-term preventive measure. Fig. 1 shows the flooding level of October 1996. The level of the -540 m drift will probably be reached in June 1997. From that point of time a general rearrangement of the ventilation has to be made because the central return air ventilation via shaft 373 will no longer be possible. This shaft has no connection with the superficia1 system of mine openings. A new return air ventilation system based on shaft 382 as return air shaft will be designed (Fig. 1). The capacity of the future mine fan system will e.g. be determined by the flow conditions in the deformation area mentioned above. For this reason investigations were carried out in a test drift 13a on the level of the Markus-Semmler-Sohle in the stretching zone of the subsidence trough for several days (Fig. 2 and 3). INFLUENCE OF THE OPERATING CONDITIONS OF THE MAIN MINE FAN ON THE RADIATION SITUATION ON SURFACE The beginning of flooding in 1991 was also the beginning of the investigations into the options of future ventilation due to the reduction of the mining area. The first analyses of radon emanation made in May 1991 showed that only about 20 % of the total radon emanation result from systematically ventilated mine workings. This means that despite of existing dams the air streams through the majority of mine openings of the deposit. The analysis of the individual radon inflows illustrates the special problematic nature of the deposit involving all possible effects on the surface. The shutdown of the main mine fan over the long period between 2.10. and 6.10.1991 showed the essential influence of the operating conditions of the fans on the whole radiation situation on surface. Fig. 4 shows the comparison of the radon concentration level and its daughters during a &apos;normal" weekend with the level during long-term shutdown in the old &apos;Casino" of the papermill of Niederschlema. This papermill as well as the &apos;Casino" are situated in the outcropping zone of the deposit section of Niederschlema. The -60 m level was driven as the upper level in this area, from which the panels were driven to about 20 m below the surface. Also below the "Casino" there is such a panel in the vein &apos;Glirck", whose roof is directly situated below the building in a depth of 39 m. Directly next to it there is the raise 2 of the panel &apos;Namenlos" in a depth of 27 m. The graphics clearly shows the immediate increase in radon concentration as well as in the daughter concentration after the shutdown of the fans. The elimination of the artificially generated negative pressure results in the reversal of the convection direction due to the natural ventilating pressure. The interface structure allows the mine airto pass through aero- dynamic connections to the surface. Neglecting the meteorological parameters the concentrations of radon and its daughters in- creased by about 350 % as compared to a &apos;normal" weekend during longterm shutdown (i.e. in 4 days). Furthermore it could be noticed that a stationary state between the radon and its daughters has not been established in the period

Jan 1, 1997

By Frank G. Horino, V. E. Hooker

INTRODUCTION The determination and use of mechanical properties of rock in engineering and rock mechanics are rapidly developing. Many of these properties are determined on intact rock specimens; thus, their application and repre¬sentation of rock mass properties may be limited. How¬ever, relative information often provides useful guidance in the solution to mine design and stability problems. Summarized in this chapter are some of the stan¬dardized techniques and procedures currently used to obtain these mechanical properties. Typical applications of the use of these properties are also presented. Stan¬dardized techniques include those advanced by the American Society of Testing and Materials (ASTM), International Society of Rock Mechanics (ISRM), US Bureau of Mines (USBM), Canadian Dept. of Mines and Technical Surveys, South African Institute of Min¬ing and Metallurgy, and other individual investigators. Information on the mechanical properties of rock and the behavior of the rock under a given system of stresses represents a necessary part of the information for rational engineering design for any given mining op¬eration. The mining method, the type and extent of sup¬port, the extraction ratio, the overall dimensions of the mine, and the orientation of the rooms and pillars are all decisions that are influenced by the mechanical prop¬erties of the ore, roof, and floor material under various stress systems and the magnitude and direction of the in situ stresses (Hooker, Bickel, and Aggson, 1972). Initial mechanical property information regarding a structure or mine property is generally obtained by two basic techniques: (1) static and dynamic property tests are conducted on intact and fractured rock specimens of exploratory drill core, and (2) dynamic properties are obtained by borehole logging techniques. When mining access becomes available, and as the mining horizon is expanded, additional information can be ob¬tained to verify preliminary mine design values. This chapter presents some of the standardized tech¬niques and equipment currently used in obtaining me¬chanical property data in the laboratory. The properties considered are: (1) uniaxial compressive strength of intact rock core specimens, (2) uniaxial compressive strength of rock cores containing planes of weakness, (3) triaxial compressive strength of intact rock core specimens, (4) triaxial compressive strength of cores with a plane of weakness, (5) Young&apos;s modulus, (6) Poisson&apos;s ratio, (7) density or apparent specific gravity, (8) modulus of rupture, (9) indirect tensile strength, and (10) creep characteristics. Where possible, an at¬tempt will be made to evaluate each property measure¬ment in relation to the problems of rock mechanics and application of results. TEST SPECIMENS The selection and care of drill core for laboratory testing require some consideration. It is recognized that laboratory-determined properties are not necessarily rep¬resentative of an in situ rock mass property. However, relative information between beds or zones of interest is still valuable information in selecting mining horizons and preliminary design criteria. To provide statistical data the number of drill core samples selected to repre¬sent each of the areas of interest should be from a mini¬mum of three to a maximum of ten test specimens. A judgment must also be made on site as to whether the recovered drill core should be wrapped and sealed in plastic to preserve moisture. On the one hand investiga¬tions of air-dried and saturated specimens have shown that moisture significantly affects the elastic properties and strengths of many rock materials (Obert, Windes, and Duvall, 1946; Colback and Wiid, 1965); on the other hand it is apparent that most core drilling is done with water which may saturate the specimen to a greater extent than in the in-situ condition. Whether or not the decision is made to retain the moisture, the core should be delivered to the laboratory as soon as possible after recovery for subsequent specimen preparation and testing. Specifications Shape: The shape of the specimens influences lab¬oratory testing in two ways: (1) time and cost of sam¬ple preparation and (2) strength of the material. Cy¬lindrical specimens of drill core are by far the least time-consuming to prepare for static or dynamic labora¬tory testing. In addition, the cylindrical shape lends it¬self to a more uniform stress distribution throughout the sample than other shapes, such as rectangles and hexa¬gons. The compressive strengths of various shapes have been studied (Grosvenor, 1963, and Price, 1960), and results indicate that the cylindrical specimens usually provide the highest strength for a given height-diameter ratio. However, reduction in strength from a cylindrical shape to a rectangular in situ pillar is not regarded as significant in relation to other considerations such as planes of weakness in a pillar or safety factors in the design process. Length-Diameter Ratio: The length-to-diameter ra¬tio, LID, has a significant effect on the compressive strength. Various recommendations have been made to use standard LID ratios ranging from 2 to 2.5 to 3 (ASTM, 1975c; ISRM, 1972). However, past work by others such as Obert, Windes, and Duvall (1946) has shown that excellent results can be obtained using LID ratios from 2 > (LID) > >/s. In selecting an LID ratio for testing, one should keep in mind the amount of material available for testing. In many instances, this may be limited. Thus, a shorter specimen such as 1: 1 LID may be necessary to provide enough test data for statistical analysis of results. Sec¬ond, it may be desirable to obtain elastic constants dur¬ing the test. This generally requires instrumentation such as linear variable differential transformers (LVDTs) or strain gages near the center of the specimen. In this case, an LID of 2.5 or 3 is desirable so that the instru

Jan 1, 1982

By W. J. Schlitt

Introduction Heap and dump leaching are low-cost metal production techniques that are gaining in popularity among gold and copper producers. However, the flow of solution in heaps and dumps has received little attention in the literature. This is unfortunate since solution flow is one of the few parameters subject to operator control. Thus, solution management may well influence both operating costs and plant performance. Costs aside, there are two important aspects of solution flow to be considered -the metallurgical and the hydrological. Some of the metallurgical factors have recently been discussed (Jackson and Ream, 1980; Schlitt, 1984; Schlitt and Nicolai, 1987). These cover solution application rates and methods, irrigation rates, leach-rest cycles, and nonvertical solution flow. According to Caldwell and Moss (1985), however, the hydrological aspects may be more significant than the metallurgical ones. In particular, a phreatic surface will form when excessive flow leads to an accumulation of water and the associated buildup of pore pressure within the rock pile. Such internal flooding can either originate at the foundation or at some other zone of very low permeability. The latter condition gives rise to an impounded volume of solution, i.e., a perched water table (see Whiting, 1985). Caldwell and Moss point out that a rising zone of saturation is probably the most common cause of dump failures. Of course, such failures have even occurred in heaps that were carefully prepared for leaching (Milligan and Engelhardt, 1984). Thus, flooded conditions and perched water tables represent an important safety consideration as well as having an impact on metallurgical performance. The following sections describe a case history in which a perched water table developed within a copper leach dump. The description includes background information, solution flow rates, and metallurgical data. Then this situation is compared to one involving normal drainage. Description of the leach system The leach dump in the case history is located at a shutdown open pit copper mine. It was built in two lifts, with the second added some 20 years after the first. There is little detailed information available on the initial lift. It was built with rail¬hauled waste and was generally less than 30 m (100 ft) high at the crest. The area available for leaching was about 120 m (400 ft) wide and more than 380 m (1250 ft) long. The dump surface was prepared for leaching by dozing ponds approximately 12 m by 12 m by 2.5 to 3.0 m deep (40 ft by 40 ft by 8 to 10 ft). The ponds were leached by flooding with barren leach solution returned from a scrap iron cementation plant. Based on mill feed at the time, the average waste grade was probably close to 0.4% Cu. The first lift leached well. It accepted high flows, which together with the waste grade, produced a rich pregnant leach solution (PLS). Old records indicate a PLS of "50 lb Cu/ 1000 gal," or about 6 g/L. As later drilling would indicate, such a high tenor led to dissolution of considerable scrap iron that was returned to the dump. The iron then hydrolyzed and settled out on the pond bottoms. In addition, the waste settled substantially so that the unleached crest was 3.0 to 4.5 m (10 to 15 ft) above the ponded area. Eventually, the leachable copper was extracted and the dump became less permeable. Thus, the PLS tenor dropped until it became uneconomical. About ten years later, mining resumed and the decision was made to add another lift to the dump. This was done without giving much thought to a subsequent leach operation. Hence a 24-m (80-ft) lift was built on top of the original dump. The surface of the latter was not prepared in any way, e.g., by leveling and/or deep ripping, prior to over-dumping. Examination of subsequent drill cuttings indicated that the new lift contained about 0.2% Cu, with chalcopyrite and chalcocite (50:50) being the predominant copper minerals. Most of the chalcocite occurred as rimming on the abundant pyrite, with the pyrite to chalcopyrite ratio estimated at 10 to 1. The use of 100- and 120-ton trucks for haulage caused some waste compaction during emplacement. In addition, the host rock itself was relatively soft, being a porphyry intrusive material that was partially altered to clay. As a result of the initial compaction and clay swelling, the rate of water percolation from the new leach ponds was slow and the ponds often contained considerable standing water. Even frequent ripping failed to provide a sustained improvement in the percolation rate. The poor surface permeability was exacerbated by hydrolysis of iron salts which settled as a layer on the pond bottoms. Partly as a result of the permeability problems, metallurgical performance was not up to expectations. These had been based on laboratory tests which showed about 20% copper solubilization in two weeks. Continued copper extraction in the tests also suggested that a substantial percentage of the copper would eventually be recovered. However, the PLS grade in the actual operation peaked briefly at about 0.48 g/L Cu (4 lb Cu/1000 gal), then declined to a range of only 0.24 to 0.36 g/L Cu (2 to 3 lb Cu/1000 gal). The poor leaching was traced to a lack of oxidation of the sulfides. There were two principal observations supporting this conclusion. First, there was no evidence of any heat being generated within the dump. As discussed elsewhere (Schlitt and Jackson, 1981; Hiskey and Schlitt, 1982), pyrite oxidation is quite exothermic and the high pyrite content of the waste should have led to an increase in the temperature of the leach solution as it percolated through the rock pile. Second, there was no sign of any natural convective air flow through the dump.

Jan 1, 1987

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 B. P. Boisen

INTRODUCTION Application Measurement of rock deformation is essential to stability and behavioral monitoring of underground openings. The difficulties of obtaining adequate sample distribution, density, and orientation can frequently be overcome by the use of single and multiple position borehole extensometers. In its simplest form, a borehole extensometer consists of an instrument head, usually anchored in the collar of a borehole, and a number of in-hole anchors, each of which is fixed in place at a known distance from the collar. As rock adjacent to the borehole deforms, the component parallel to the borehole direction causes a distance change between adjacent in-hole anchors, as well as a distance change between individual anchors and the instrument head. These changes are measured by means of a mechanical link (usually wire or rod) extending to the instrument head from each anchor. The resulting data can be used to compute the distribution, magnitude, rate, and acceleration of deformation in the material intersected by the extensometer&apos;s borehole. Wire Vs. Rod Extensometers Since the early 1960s myriad varieties of borehole extensometers have become available through the combination of various special features of two basic instruments, the wire and rod versions of borehole extensometers. Recent publications have recommended shallow depth limitations for wire extensometers, and favor rod extensometers for a broad range of borehole inclinations and depths. There are three main reasons for using a deep wire system: its cost effectiveness; the fact that tremendous depth means tremendous rod weight which will cause the rod to buckle or rub against the borehole or casing wall; and users&apos; attitude and proper understanding of how to install and apply a wire-type extensometer. Here experience plays a big role. Over the years, many have learned the hard way that untensioned rod extensometers are simple and inherently inexpensive, but limited to vertical up holes where sensitivity is not influenced by friction between the rod elements and the borehole wall or protective conduit, the inertia of the rod elements, or the direction of the movement to be measured. In deep, horizontal, or inclined holes, and in instances where compressive as well as tensile strains or displacements are anticipated, tensioned rod or wire systems are preferred. System Components Borehole extensometers are made up of at least the following five principal components: the instrument head; the collar anchor assembly; the downhole system of anchors and rods or wires; the signal cable if the head is equipped for electronic readout; and the readout apparatus, whether mechanical or electronic. These components are shown in Fig. 1. Borehole Requirements Most extensometers are designed for installation in nominal 75 to 57-mm (3.0 to 2.25-in.) diam drill holes. Either a diamond drill or a percussion drill hole is satisfactory, provided the variation in drill-hole diameter is within the tolerance of the standard anchor design. The basic system may be modified for other hole diameters by changing the dimensions of the collar anchor and downhole anchor assemblies. Some instrument heads and collar anchor assemblies require enlarged borehole collars, though usually not over 150 mm (6 in.) for a depth of 1 m (3 ft). It is desirable, although not essential, that holes be free-draining. Holes must be clean and free of debris before installation. Depth/Orientation of Extensometer Boreholes Extensometers are designed to measure axial strain components in the instrumented drill hole. As the drill hole is axially deformed, the distances between individual downhole anchors and the instrument head change, as do the distances between adjacent anchors. The distance changes measured by a multiple position borehole extensometer, for example, are reduced to functions of time, hole depth, anchor position, or other parameters of significance in terms of the structure under investigation.

Jan 1, 1982

By Agne Rustan

INTRODUCTION This work is a continuation of the preparatory study by Rustan and Stöckel 1979 (3) and (4) of the possibilities to introduce automatic mine ventilation in Swedish mines and construction works. The preparatory study has been financed by the Swedish Board of Technical Development. In the study it was recommended that an automatic step by step control of the fans near the working place due to diesel exhaust emission would be the best solution. A more advanced alternative would be to regulate the fresh air continuously in proportion to the contamination. The latter alternative has been studied by Boliden Mineral Co in the Udden mine. In this study it was found that the air supply could be regulated continuously but it was very important where the sensors feeling the contamination were placed. An intense study of the environment underground is now being planned by our division together with specialists from the National Board of Health and Safety, The Swedish Work Environment Fund, LKAB and the local Governmental Inspection of the Workers Safety. Different gases, radon, dust, airspeed, moisture temperature and pressure is planned to be measured continuously and correlation between the parameters is going to be established to decide the best measurement parameters for regulating the system. The best situation of sensors is also going to be studied. Despite this work hasn&apos;t been done yet this paper covers some of the problems which could be encounted with continuous regulation. In this report the environment has been studied during the test with step by step regulation of fans by radio. Diesel vehicles are the largest polluters under ground in Swedish mines and we have therefore suggested that the large diesel vehicles should regulate the fans. The goal of this study has therefore been to find how active diesel vehicles should be able to start and stop gallery fans at both intake and return air. In this paper different methods for activating fans are discussed. Then the best methods are chosen and tested in the mine. In this study we have found that changes of temperature is a cheap and effective method to start and stop cross cut fans in sublevel caving and radio signals seem to be an attractive system for start and stop of gallery fans. This work has been done in close collaboration between LKAB in Kiruna and the University of Luleå - Division of Electronics and Division of Mining and Rock Excavation. A reference group for the project has been established with ventilation experts from the mines, representatives for the local mine worker&apos;s unions and authorities from the local Governmental Inspection of the Workers Safety. VENTILATION SYSTEM STUDIED In the project the ventilation system at LKAB in Kiruna has been studied. The mining method used is sublevel caving with 9-13 m slice height. The air is heated winter time over ground by hot water heat exchanger. The main fans are pushing air down to a main distribution level along the 4000 m long and about 100 m wide orebody. From the distribution level, air is taken through shafts down to the sublevels, where it is pushed through the levels by local gallery fans on both intake and return side, see fig 1. The air ( ~ 8 m3/s) is moving freely through the gallery head. The cross cuts are ventilated with smaller fans, diameter 500 mm, and flexible duct, see fig 2. The aim of this study has been to study how to start and stop gallery- and cross fans and examine the environment when the fans are stopped . In Sweden there are no regulations telling that a minimum of air amount must be supplied. Instead the hygienic limits must be fullfilled.

Jan 1, 1981

By Sidney A. Hiscock

INTRODUCTION Lead and zinc are today rightly regarded as sister metals. Historically, however, they differ markedly with lead being known and in general use as the metal since 3000 BC in several countries which at that time might have been described as mature economies. Zinc was not isolated and recognised as an industrial metal with distinct and valuable properties until about 1700-1800 AD (although it had, of course, been utilised as a constituent of brass for some 2000 years previously). c were well established as industrial metals by the beginning of the 20th century. The present paper briefly reviews the overall growth in world consumption* and changes in the three main areas which today represent the mature economies or industrialised areas, i. e. ,Europe, the United States and Japan. Trends in overall consumption are determined by the demands for the individual uses for lead and zinc. The changes in patterns of use in the industrialised areas - individually and collectively - and reasons for the changes are also considered. LONG TERM TRENDS - 1900-1984 The World Picture Since the beginning of the 20th century the world consumption of lead (3.9 million tonnes in 1984) has grown almost fourfold and that of zinc (4.7 million tonnes in 1984) is almost eleven times greater than it was in 1900. The long-term trend has been one of continuing expansion for both. metals although the pattern has been severely disturbed at certain periods, for example in the years immediately following the two world wars and at times of marked recessions in industrialised areas. Apart from such major setbacks, new levels of consumption have been established regularly every few years. Recently, however, consumption has not really grown and with only a slow and partial recovery in world industrial activity at present, there is no indication that the peak consumptions of 4.2 million tonnes of lead, set in 1979, and 4.8 million tonnes of zinc (1973) will be exceeded in the near future. The growth in world lead and zinc consumption since 1900 is shown in Figure 1 (which includes changes in copper and aluminium for comparison). Overall growth over the period 1900-1984 has been about two percent a year for lead and about three percent for zinc (annual growth for copper has also been about three percent). Increases in tonnage consumption and growth rates by decade for lead and zinc are summarised in Table I. However, the growth rates by decade sometimes conceal very large individual annual increases and decreases. For example, in 1975 lead consumption fell by 13 percent and zinc consumption by 22 percent compared to 1974. In 1976 consumption of lead and zinc rose by 12 percent and 18 percent respectively. The years when consumption &apos;landmarks&apos; (ie one, two, three and four million tonnes) were first reached are shown on Table 2. Clearly, the pattern of shorter periods being required to attain each extra million tonnes of consumption was broken in the 1970&apos;s. The consumption of zinc overtook that of lead for the first time in 1940, and since 1946 has always been at a higher level. Trends in industrialised areas The two key industrialised areas at the beginning of the 20th century were Europe and the United States which between them accounted for some 95 percent of world lead consumption and 98 percent of zinc consumption. Figures 2 and 3 show the growth in lead and zinc consumption, and Figures 4 and 5 the percentage shares for each area since 1900. Europe has usually been the major user of lead except for some years in the 1920fs, most of the 40&apos;s and early 50&apos;s when consumption in the United States was larger. In 1900, Europe accounted for about 64 percent of world lead consumption and the United States just over 30 percent. Currently, Europe takes about 40 percent of world consumption and the

Jan 1, 1986

By Niels Jonassen

INTRODUCTION The occurrence of the radioactive element radium, [2gg~a], often in trace amounts, in the crust of the earth gives rise to a flux of the daughter product radon, [&apos;ii~n], across the earth-air interface. This so-called exhalation process has been studied by many investigators [KO 45, Kra 64, Pe 66, Wi 741] over several decades. In the earlier investigations the exhalation process was studied in a purely geophysical context, but over the last 20 years it has been realized, that under certain conditions the exhalation may be intense enough to also have radiological implications. This is for example the case where relatively radium-rich wastes or by-products, like uranium mill tailings and phospho-gypsum, are stored in larger quantities. In order to be able to predict the radiological impact on the environnent it is necessary to know the rate of exhalation of radon from the surface of the exhaling material. Usually the determination of these exhalation rates has been done by measurements in situ directly upon the accumulated material [~le 74, Ha 791. Although such field measurements have the obvious merit of dealing directly with the full scale problem, the results are often lacking in generality, making predictions for other situations difficult or even impossible. The present paper describes some results of an investigation where it is attempted from laboratory measurements on samples of limited size to obtain material parameters, which are suited for extrapolating and predicting the radiational behaviour of larger accumulations of the material. THEORY Let us consider a sample of a radon exhaling material enclosed in a container with a dead space volume, Vd, and initially containing radon free air. If the total exhalation rate (number of atoms per unit time) is c, the radon concentration (number of atoms per unit volume) N will initially grow with the time t according to the equation [ ] where [R] is the decay constant of radon. As the concentration (or activity) in the container grows, the net exhalation rate will slow down because of back diffusion, and, furthermore, radon atoms may be removed from the container, not only by radioactive decay, but also by diffusion leakage [JO 81 ]. If, however,only the initial. part of the activity growth curve is considered, the total exhalation rate, c, is given by [c = r•Vd(2) where r is the initial activity growth rate in the container. Let the container be cylindrical with an inner cross-section S and filled to a height L with a radon exhaling material with the specific mass p, the porosity E, the radon pore production rate f and the diffusion length i. When the container is left open,i.e. the radon exhalation takes place into a virtually radon free room, the radon area exhalation rate, E, can be written [Jo 80] E = Ef?t tanh L.R (3) assuming the exhalation only takes place in one direction, towards the surface of the material. It appears that for values of L > 2•ithe area exhalation rate becomes E a = Efi (4) independently of the sample dimensions and the total exhalation rate c is then c = E •S (5) For L ;~ 0.5•i, tanh P _ k and equation (3) may be written E ~ EfL (6) then c = EfL•S = p•m= Em •m (7) where m is the mass of the sample and Em = P (8) is the mass exhalation rate. It thus appears that the radon exhalation rate from a given amount of material may often be conveniently expressed by c = E •Sfor L z 2•I (5) a orc = Em •mfor L0.5•i (7) where S is the total exhaling area and m the total exhaling mass, E and E the area and mass exhalation rates respectively defined by equations (4) and (8).]

Jan 1, 1981

By Justus B. Deen, Randy Peterson

INTRODUCTION The General Chemical Soda Ash Operation located near Green River, Wyoming produces about 4.5 Mt of trona per year. In July 1989, Mine Ventilation Services Inc. performed a ventilation survey of this complex trona mine which had thirteen active panels spread over 36 square kilometers. The ventilation survey and subsequent study showed that the intake shaft air velocities were at their practical limits (Wallace and Rogers, 1987). The need for additional intake capacity was immediate. The ventilation network modeling delineated the best location for this new intake shaft as it applied to the 5 year ventilation plan. Unfortunately, this location, 6.5 kilometers from the soda ash processing plant, would make heating this intake air very costly. Presently, the intake air is heated by heat exchangers which use waste heat from the plant. Air heating during the winter months is necessary for miner comfort and to prevent potential freezing of water lines in the mine. It was decided to examine the feasibility of heating the air geothermally by coursing it through an old production panel. This method of air heating has been used in this region with good results at the Stauffer Chemical Operation (now Rhone-Poulenc) (Moore, 1985). In 1990, a geothermal heating study examined four different production shaft/panel configurations and a workable design was found. The operators at the General Chemical Soda Ash Operations acted on this design and began raise boring a ventilation shaft in the Spring of 1991. This paper describes the ventilation system, briefly, and the geothermal heating studies performed. Findings of the 1990 geothermal studies are then compared to field data acquired in the Winter of 1992. THE VENTILATION SYSTEM The ventilation system at General Chemical had four shafts serving the mining horizon in July 1989; three of these shafts are located at the Northwest end of the mine. The active work areas had progressed to the South over the past years and supplying ventilation has become increasingly difficult. The 6.1 meter diameter #3 Production shaft and the men and materials compartment of the 6.1 meter diameter #2 Split Shaft had reached their air velocity limits. Conversely, the 4.6 meter diameter #5 Exhaust Shaft was under utilized. Ventilation studies showed the conversion of the 3.7 meter diameter #1 Exhaust Shaft to an intake would provide adequate intake capacity for two years. The next option to improve ventilation was the addition of an intake shaft at the southeast part of the mine at the intersection of H Mains South(H-M-S)and J Mains East(J-M-E) (See Figure 1). A service shaft has been planned for the extreme Southeast end in the late 1990&apos;s. Ventilation network analyses showed this interim intake shaft would have to be at least 2.5 meters in diameter to postpone the service shaft construction beyond 1995. The greater the size of the borehole, the further the construction could be postponed. In October 1991, General Chemical commissioned the 4.57 meter diameter #4 Ventilation Shaft near the junction of H-M-S and J-M-E as shown in Figure 1. The L95 panel would be used for air heating as described in the geothermal heating section to follow. GEOTHERMAL STUDIES Geothermal heating studies were performed by MVS Inc. in May 1990. Using a modified code of CLIMSIM, an underground mine climatic simulation model, several different shaft/panel configurations were examined to see which would provide adequate air heating during the winter months. CLIMSIM simulates heat flow into a single underground airway. The program was used to evaluate the heat flow to and from the airway surface and to calculate the resulting change in dry bulb temperature of the ventilating air. CLIMSIM uses inlet air conditions, airway characteristics, and rock thermal properties input by the user to predict the variation of psychrometric and thermodynamic parameters along an underground airway.

Jan 1, 1993

At present the feasibility of developing underground facilities for the storage of a number of products including toxic waste, oil, nuclear waste and low pressure gas is under investigation in e number of countries. A possible method of storage is in large excavations, Preliminary ventilation design work has identified that the climatic conditions encountered within these large excavations are a limiting ventilation pollutant within the ventilation system. This paper examines the design of a variety of sizes of excavation from a ventilation viewpoint. The main phases of an excavations lifetime can be divided into four main phases, these being construction, civil and mechanical fitting, storage, full and sealed. The paper details ventilation design options for the first two phases of e vaults lifetime, giving climatic simulations to assess the likely climatic conditions within the excavations for e number of facility depths. The options for climatic control are given and analyzed by means of climatic simulations techniques. l. INTRODUCTION The feasibility of developing underground facilities for the storage of a number of products is under investigation at present in a number of countries. This paper details a possible design for such a facility where the product is to be stored in large underground excavations. A facility layout has been designed with the aim of allowing concurrent construction and storage operations. Thus from a ventilation perspective the facility consists of two circuits , a construction ventilation circuit and a product ventilation circuit both of which are served by an intake and a return surface connection The construction and storage circuits are considered as independent of each other , any interaction in the form of leakage must be from the construction to the storage circuit. From the facility design a ventilation network was developed to assess the ventilation requirements of the facility. In analyzing the pollutant loads for the construction circuit two pollutants were recognised to be critical to the ventilation design , these being diesel exhaust fumes and heat end humidity. In this paper ventilation design options for the construction and civil and mechanical fitting phases of the excavations are analyzed by means of climatic simulation and options for climatic control within the vaults are outlined . 2. FACILITY DESIGN A schematic diagram of a facility design is illustrated in Figure 1. There would be two distinct ventilation circuits , one to service the construction operations and the second to serve the storage operations which occur concurrently with construction , the circuits are separated by ventilation bulkheads . Access to the facility for intake air is provided by two drifts, one to serve each circuit. Two shafts act as returns to surface. The facility consists of two distinct zones for the storage of two types of product. One type is to be stored in smell linear excavations. The second zone is for the storage of a second product in larger excavations. The maximum cross-sectional area and perimeter for the smell linear excavations are 456.25 m2 and 83.2 m, while for the larger excavations they are 440.1 m2 and 86.4 m. Excavation construction consists of three phases: 1. Initial pilot tunnel development between main intake and return to provide through flow ventilation. 2. Enlargement to produce the required finished cross-section. 3. Civil and mechanical engineering work to fit out the excavation ready for storage. It is proposed to excavate the excavations using drill and blest methods utilising diesel equipment. Spoil is to be transported to surface by a conveyor belt system. At any one time the maximum number of working places requiring ventilation will consist of ; 3 storage sites, 2 excavations under construction, 2 excavations undergoing civil and mechanical fitting. [ ]

Jan 1, 1993

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

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

Jan 1, 1981

By Donna L. Boreck

Introduction Coal has been mined in the western United States since the mid-1800s. The resource is abundant in the West. Despite recent economic conditions, it remains an important part of the nation&apos;s energy warehouse. Many of the seams that comprise the western resource base are thick, some exceeding 61 m (200 ft). Multiple, closely-spaced seams or thick coals that split into two or more minable seams are also common. Historically, subsurface development recovered from less than 15% to 50% of these deposits (Matson and White, 1975). This resulted in a substantial loss of inplace reserve due to the limited height that could be mined using the available technology. In an effort to increase future recovery from such deposits, the US Bureau of Mines analyzed the potential for incorporating three thick-seam mining methods into western coal mines. These methods are high-face longwall, multislice longwall, and longwall caving. They were developed in Europe and are being used to mine thick coal seams in Europe and Asia. One aspect that will determine their success in the US is the geology of the individual deposit; namely, are the methods adaptable to the geologic conditions characteristic of western coal? This paper analyzes some of the geologic factors that would affect development of thick or closely-spaced seams using these mining methods. Occurrence of thick and closely-spaced seams Since mining conditions and methods differ around the world, a thick seam is often defined as a seam whose total thickness cannot be extracted using the available equipment. In the western US, the maximum extractable mining height is about 4.3 m (14 ft). For this research, a thick seam was defined as a minable seam 4.6 m (15 ft) thick or more. Since one of the seams is usually lost using standard mining methods, closely-spaced or split seams were also addressed. These were defined as two minable seams 1.5 m (5 ft) thick or more separated by 9 m (30 ft) of interburden. Thick or closely-spaced coal seams are common in the West. Their importance has not been documented due, in part, to the methods used to measure resources. In the past, all reserves greater than 1.1 m (42 in.) thick for bituminous seams and greater than 3 m (10 ft) thick for subbituminous seams were grouped together and were differentiated only by rank. Further differentiation by seam thickness or minability was not done. For this reason, reserve figures for thick or split seams are difficult to determine. By analyzing available published resource data, Ferm and Muthig (1982) addressed thick seam resources. They estimated a resource of 90 Gt (99 billion st) of coal (16% of the Western resource base) in seams 4.6 (15 ft) or greater at less than 610 m (2000 ft) overburden. Yet, a later study of Pierce and others (1982) on one thick seam deposit in the Powder River Basin estimated a resource of 103 Gt (113 billion st). Pierce (1986) reported that the overburden reached a maximum of 731 m (2400 ft). The resources in this single deposit exceeded the 90 Gt (99 billion st) total projected for the entire western US by 12.7 Gt (14 billion st). Due to such differences, this author chose to look at reported occurrences of thick or closely-spaced minable seams in place of resource figures, to verify the need to research thick seam mining in the West. An occurrence was defined as an isolated site where a thick seam or two or more closely-spaced seams were reported. The occurrences were derived from published material - mostly mine descriptions, measured sections, coal exploration logs, and oil and gas logs. Due to time constraints, the reserves associated with each occurrence were not determined. All data points were random. The analysis was limited to occurrences in Colorado, Wyoming, and Utah. Although source material available to the author was limited to general coal summa-

Jan 3, 1988

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

Jan 7, 1980

By Patzen Wu, Chao-Ming Tsai, Shaw-Chii Wu, Yuan Lin

INTRODUCTION The development of nuclear industry in Taiwan requires an increasing supply of uranium. Uranium recovery from phosphate ores, becomes an important item of the R & D programs. Many countries have already started such a program. USA has set up an industrial-scale production plant; South Africa, Israel, Morocco and other countries have also engaged in the development of extraction technology of uranium from phosphate ores and started setting up the production plant. In Taiwan, a feasibility study of establishing the production plant was started in 1977. A pilot plant was completed, tested, and operated in 1979. The main source material. - phosphate ores is obtained from China Phosphate Company. The annual yield of the pilot plant has been planned to be one ton of ammonium diuranate (ADU). The Health Physics Division of Institute of Nuclear Energy Research (INER), Atomic Energy Council of Republic of China (ROC AEC), is responsible for the radiation safety control of the plant. The purpose of the radiation safety control is to collect radiation safety information of the plant of this type, to evaluate possible radiation hazards to the personnel, and to analyze the radiation safety of the immediate environment adjacent to the plant. Furthermore, the information and experience obtained are used as the basis for applying radiation safety control to the production plant. The main emphases of the radiation safety control are therefore the arrangement of radiation safety control areas and the collection and evaluation of the health physics data. This report includes the past one and half years&apos; experiences of the radiation safety control of the pilot plant, the planning of the production plant with an annual yield of 10 ton ADU, and considerations given in improving the radiation safety control of the production plant. PILOT PLANT AND ITS OPERATION PROCESS The pilot plant occupies an area of 350 m2 and is a two-story rectangular building. Exits are provided by a front and a rear door. An additional small door at the central right side of the building is connected to the maintenance area located outside the main building. The interior of the building is divided into interconnected offices, laboratories and process area (Figs. 1 and 2). A stairway to the process area of the second floor is located in the central part of the first floor. The second floor offices are accessible by the additional front and rear stairways. The main process area of the first floor is furnished with plastic floor, and the second floor with poriferous iron-grid floor. The building is not furnished with centralized air-conditioning system. Individual air conditioners are, however, available to some laboratories. Because the back door of the building is frequently open, the product storage room is provided with a blower to maintain good ventilation. Natural uranium contents of phosphoric acid solution vary in the range from 0.07 to 0.11 kg/m3. The process consists mainly of the uranium concentration cycle and the purification cycle. A flow diagram of the process is shown in Fig. 3. CONSIDERATIONS GIVEN FOR RADIATION SAFETY CONTROL AND THE ESTABLISHMENT OF RADIATION CONTROLLED AREAS The radiation safety control of the pilot plant is carried out on the basis of the considerations listed below:1,2 1) Since the wet process is adopted, and the flow of the radioactive material is confined to closed pipes and containers, air contamination is therefore not too much of concern. 2) The concentrations of radioactive material of the whole process are mostly less than 10 ppm. An exception is the final ADU product after precipitation and centrifugal separation, which contains up to 50% natural uranium. It is therefore necessary to apply radiation safety control to the process area where

Jan 1, 1981

By Frank W. Schmechel

The subsidence monitoring program at the Old Ben No. 24 Mine was planned in 1976 as part of the on- going demonstration of longwall mining. The construction and installation of the network of surface subsidence monuments was started in Play of 1976 and completed four weeks later after which an initial reference survey was made of all surface monuments. Mining of the first longwall panel started in September of 1976 and was completed in May of 1977. Mining of the second panel started in August of 1977 and is continuing at this time. The Automatic Data Acquisition System (ADAS), which is the subject of this paper, was installed over the second panel on March 29, 1978 near the panel center (Figure 1) and approximately 104 meters (370 feet) in front of the face position at that time. Mine No. 24 is located approximately 5 kilometers (3 miles) Northwest of Benton, Franklin County, Illinois. Coal is being mined from the Illinois (Herrin) No. 6 seam which is at a depth of about 189 meters (620 feet) at the site. The Herrin No. 6 is up to 2.7 meters (9 feet) thick in the mine but a constant thickness of 2.1 meters (7 feet) is being mined by the demonstration longwall with both roof and floor coal being left for stability. Although the seam is flat laying at the mine, it dips north- ward regionally. Topographically, the surface over the demonstration panels is essentially flat and low laying with several areas of standing water. The Rend Lake Dam (United States Army Corps of Engineers) is about 305 meters (1000 feet) north of the demonstration site.

Jan 1, 1979

By C. D. Atkinson

The past few years have seen a much greater concern about air pollution in this country. This concern has led to increased legislation restricting the discharge of many substances such as sulfur dioxide and particulate matter to the atmosphere. In the future, it is anticipated the present permissible emission levels will be lowered. Current and impending restrictions on the emission of sulfur dioxide have brought out the fact that there is a shortage of available low-sulfur fuel for power plants and industrial use which promises to have an economic impact on the nation&apos;s energy picture (1) The Pittsburg & Midway Coal Mining Co., under the sponsorship of the Office of Coal Research, has developed a process to produce a fuel capable of meeting the energy requirements of this nation (2) The process is called "The Pittsburg & Midway Solvent Refined Coal Process". The principle product from the process is Solvent Refined Coal which is an ashless, low-sulfur fuel with 16,000 BTU per pound regardless of the rank coal from which it is produced. The product can be burned as either a solid or a liquid fuel. The process appears to be economically more attractive than other methods of alleviating the pollution problems of this nation. In addition to reducing the air pollution caused by stack gases, the product produced is potentially useful as a source of electrode coke and electrode binder pitch, as a gas turbine fuel, in the production of metallurigical coke and pavement asphalt, and as the first step in the overall process to produce distillate fuels and pipeline gas (3).

Jan 1, 1973

By B. N. Everett

The growing role of Colombia as a major supplier of internationally traded steam coal has been widely discussed in the industry. Although Colombian coal has been produced and exported in small quantities for a number of years, Colombia&apos;s position as a major coal exporter began in earnest in 1985 with the first commercial exports from the Cerrejon North Block Project. By way of background: ? The mine is located on the Guajira peninsula in Northeast Colombia, about 90 miles from the coast, as shown in Figure 1. ? In addition to the mine itself, the project involves construction of a dedicated railroad and port, which form a modern, efficient and integrated system. ? The Project is embodied in a 50/50 Association Contract between International Colombia Resources (Intercor), a wholly owned subsidiary of Exxon, and Carbocol, the Colombian state coal company. Intercor and Carbocol share investment and operating expenses equally and Intercor has the additional responsibility of managing operations on behalf of both partners. Each partner is entitled to 50% of the coal produced, and each markets its share separately. ? The project is a multi-seam, truck and shovel operation, based on large reserves (2-3 billion tonnes&apos;) of high quality bituminous coal. ? Exports began in the first quarter of 1985 at modest levels through what we call Early Coal Facilities, designed and constructed apart from the permanent coal handling system. ? Production will increase rapidly in each of the next few years, reaching the design capacity of 15 Mt per year in 1939. Based on available resources, however, the resource will support expansion, and continued growth n production into the 1990s appears feasible.

Jan 1, 1986