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Collaborating Authors
The 25th U.S. Symposium on Rock Mechanics (USRMS)
ABSTRACT ABSTRACT In this paper the imminent relationship between rock burst and its phenomenon with normal rock pressure and its appearance, as well as the factors of one changing into another have been analyzed. Based on this, the necessary and sufficient appearing factors of rock bursts have been put forward, and the term "burst pressure" has been defined. According to the statistics data of tremendous rock bursts which appear in collieries in Hunan Province, the features of high frequency, great strength and shallow depth of the beginning burst have been found. There are some special structures in the coal seams, which are closely related with the burst phenomenon. A hypothesis of "star structure" has been suggested for the appearing mechanism of bursts. Inside and around this star structure the distribution rule of burst pressure and rock strength has been researched depending upon the theory of the limited stress field. A new description of the appearing causes of bursts has been raised. THE ESSENTIAL FEATURES OF ROCK BURSTS IN HUNAN COLLIERIES In Hunan Collieries, rushing phenomenons appear almost by way of bursting. Coal, rock, gas and water all rush out, especially the outburst of coal and gas at Ma-Tian Colliery in April 1955 to the end of 1980, 3,421 bursts had come into being during this period involving 30 mines. A great many of the bursts threatened the mine safety and production seriously as well as the measures taken to prevent them were not very efficient. Miners in China usually called them "mine cancers." The essential features of bursts in Huanan are as follows: Fast advancing speed and high appearing frequency. From Table 1 and Fig. 1, it is evident that from 1955 to 1967 the increasing speed of bursts was rather slow during these 13 years. From 1968 to 1980 it was very fast during these years and reached a peak in 1979. One hundred eight burst appeared during the first 13 years and 2793 during the second 13 years. Otherwise, in the second 13 years there were 15.5 times as many as in the first. It is very clear that the bursting frequency rises with developed mines and mining depth. The statistics data of 1980 tells us that the average bursting time per 10,000 tons of coal extracted from the mine reached 92. The possibility and strength of bursts increases as the mining depth increases. From Table 2 and Fig. 2, the possibility ยฟ, maximum Qmax and average strength Q vary as the increasing of mining depth. from 80 to 150 m, the possibility ยฟ evidently rises with the increase of mining depth. Beyond 150 m the curve ยฟ drops because the drivaging length was not long enough. The maximum strength Qmax and average strength Q rise as the mining depth increases. Shallow depth of the beginning burst. From Table 3 the beginning depth of bursts did not go beyond 100 m in 17 mines until 1981 while the static pressure of overburden weight is not beyond 25 kg/cm , and the gas pressure is 2.1 kg/cm to 34 kg/cm .
ABSTRACT ABSTRACT Underground mining will disturb the natural equilibrium of the rock mass, causing significant stress redistributions in the vicinity of the excavation with corresponding horizontal and vertical displacements. Subsidence of the ground occurs when these displacements propagate from the mine opening, through the overlying strata, to the surface. Such ground movements may take either of two general forms: a sudden and sometimes violent collapse, or a more gradual and progressive ground movement. The former, characterized by localized failure of roof or pillar in old, shallow room and pillar workings, is rather unpredictable and difficult to analyze. As a result, most control measures are directed toward the latter type of movement, which is associated with the fracture and flow of the strata overlying longwall and high-extraction room and pillar systems, resulting in horizontal and vertical strains in the affected area. Such ground movements can often cause considerable surface disturbances ranging from simple land settlement to severe structural damage. Due to the increasing incidence of this phenomenon in more populated areas and the rising awareness of the problem, it has become necessary to establish a balance between the development of mineral resources and the protection of the environment and society. This paper presents a summary of the major findings derived from a substantial research effort on mining subsidence carried out, during the past few years, in the southern Appalachian coalfield. The research program encompassed the following major tasks: identification of the most significant factors influencing ground movements above mined openings, collection of subsidence case studies for the southern Appalachian coalfield, development of subsidence and strain prediction techniques for that region and initiation of a subsidence monitoring program in southwestern Virginia. SUBSIDENCE PREDICTION Due to the lack of subsidence information for the eastern U.S. coalfields, the preliminary establishment of a data bank on surface settlement was a prerequisite before regional subsidence characteristics could be determined. Since the number of parameters involved in describing ground movements above longwall mining seemed to be less than those necessary for the characterization of room and pillar subsidence, the collection process was first directed toward the compilation of longwall case studies. All relevant published information was collected and numerous coal companies were contacted to contribute any unpublished information that might be of interest to this study. This data collection phase generated 34 longwall case studies from the Appalachian region. The more important parameters involved in subsidence modeling are given in Table 1 for these case histories. Through analytical and statistical investigation of this data, the significant subsidence relationships were determined, including the dependence of the angle of draw, the maximum subsidence factor and the location of the inflection point on the width-to-depth ratio. In addition, a correlation between the maximum subsidence and the geology of the superincumbent strata was developed. From these relationships, the dependence of subsidence on the panel geometry and amount of hardrock (sandstone and limestone) in the overburden was formulated (Figure 1). Once these subsidence relationships were established, the development of a longwall subsidence prediction model was attempted. However, before an empirical prediction model could be developed for the Appalachian region, it was necessary to determine the characteristic subsidence profiles.
- Geology > Rock Type > Sedimentary Rock > Organic-Rich Rock > Coal (1.00)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Structural Geology > Tectonics > Compressional Tectonics > Fold and Thrust Belt (0.88)
Excavation Through Highly Fractured And Sheared Host Rock - Pacheco Pumping Plant
Sinha, R.S. (U.S. Department of the Interior, Bureau of Reclamation) | Carlson, P.F. (U.S. Department of the Interior, Bureau of Reclamation) | Briggs, G.O. (U.S. Department of the Interior, Bureau of Reclamation) | Bartlett, S.R. (U.S. Department of the Interior, Bureau of Reclamation)
ABSTRACT ABSTRACT A properly designed pre- and post-grouting program aided with coherent construction sequence and adoption of flexible design approach provide a practical combination to solve excavation problems a hostile rock environment: intensely fractured and sheared. The Pacheco Pumping Plant demonstrates the utility of an engineered grouting program. INTRODUCTION In a highly fractured and sheared rock environment, the problems associated with the high potentials of caving, squeezing, and water inflow may attain a formidable dimension. Under normal circumstances, prudent planning, design, construction, and maintenance criteria will dictate that a rock excavation site in a hostile rock environment should be avoided. However, there are situations where such a compliance cannot be made. Pacheco Pumping Plant is an example of such a situation. In order to keep the project costs and time within the realm of practicality, the rock excavations for this facility were located, out of necessity, in a demanding host rock environment. The USBR expects a successfully completed project by the end of 1985. Described in this paper is the experience of the USBR (U.S. Bureau of Reclamation) in building the underground works for the Pacheco Pumping P1 ant. THE PROJECT Pacheco Pumping Plant (fig. 1), located approximately 32 km (20 miles) east of Gilroy, California, will lift water some 76 m (250 feet) from San Luis Reservoir via existing Pacheco Tunnel No. 1 and will discharge it into already constructed Pacheco Tunnel No. 2. The pumping plant which includes a battery of 12 pumping units, each suspended about 76 m (250 feet) vertically, will be commissioned to pump 13.6 m/s (480 ft/s) of water from the pump chamber below. The design requires the suspension of the pumping units from the floor of the pumping plant building on the surface through individual pump intake shafts excavated in the Franciscan complex. The pump intake shafts, designed to be steel lined throughout, will be 1.52 m (5 feet) finished internal diameter and will accommodate pump installation and pump bowl walk around resulting from the pump operations. The underground, cylindrical pump chamber is to be a continuation of Pacheco Tunnel No. 1 and is to be 6.1 m (20 feet) in finished diameter and 93.3 m (306 feet) long. The chamber will be connected to the ground surface by means of a surge shaft, 4.6-m (15-foot) inside diameter, and a battery of 12 pump intake shafts. The pumping plant, when completed, will supply water for irrigation, industrial', and municipal uses to Santa Clara, San Benito, Monterey, and Santa Cruz counties in California. GEOLOGY Pacheco Pumping Plant is located in the central portion of the Diablo Range within the Coast Range Province of central California, which is characterized by northwest-trending mountain ranges and structural valleys. The Diablo Range is a large, faulted, crude, anticlinal arch 250 km (155 miles) long and 50 km (31 miles) wide. The range contains a core of Franciscan complex rocks in fault contact with flanking clastic sedimentary rocks of the Cretaceous Great Valley sequence. Major active faults in the area make this part of California one of the most active seismic regions in the United States.
- Geology > Structural Geology > Tectonics (0.89)
- Geology > Rock Type > Sedimentary Rock (0.54)
ABSTRACT ABSTRACT Specimens of copper-bearing quartz monzonite were subjected to a plane shock wave simulating high compressional stresses in the proximity of a borehole wall. Fragmentation was studied as a function of stress levels between 1.6 GPa (23 x 10 psi) and 7.3 GPa(105.8x10 psi) and pulse durations ranging from 1.1 ยตs to 6.9 ยตs. Both explosive pressure and pulse duration were shown to have a strong effect on particle size distribution. While the effect of pressure is well recognized, the effect of pulse duration indicates that ultimate fragmentation increases with the time of application of a plane wave pulse. At high pulse durations, degree of fragmentation is more sensitive to changes in pressure than at low pulse durations. INTRODUCTION Recent interests in improving the efficiency in the comminution of ore minerals have led to studies involving the characterization of fracturing from explosive loading. Energy consumption in the mechanical crushing and grinding of ores for beneficiation is less than 1% efficient (Committee on Comminution and Energy Consumption, 1981). In 1978, electrical energy consumed to crush and grind copper ores was 18 times the energy consumed for explosive fragmentation. The increase in surface area per kilowatt-hour equivalent of explosives consumed is not well understood. Conventional rock breakage using explosives in cylindrical boreholes is performed for handling ore during mine operations; little attention is given during blasting to the amount of particle size reduction for processing. Fine crushing, however, occurs within a limited region surrounding the borehole. There is considerable interest to extend this zone of finely crushed rock during blasting. The mechanisms of rock breakage using multiple cylindrical charges are extremely complex. A number of experiments are being conducted in an attempt to understand the roles of stress waves, delayed gas pressures and their interaction with reflecting free surfaces on fragmentation. In the first stage of this investigation, the effects of the shock wave which produces the "rubblized" region in the proximity of the borehole was studied. In order to allow independent variation of the peak pressure and pulse duration, an impact technique utilizing a flyer plate was used. Ten experimental explosive events were conducted to determine the effects of compressive explosive pressure and pulse duration on fragmentation for copper porphyry in the absence of reflected waves. Nominal design pressures ranged from 1.2 to 8.0 GPa (14.5 to 116 x 10 psi) while pulse durations ranged from 1 to 6 ยตs. The axial loading is thought to simulate the high compression forces near the borehole wall during blasting. Rock specimens used for shock experiments are altered quartz monzonite porphyries consisting of quartz, orthoclase, and abundant sericite. Grain sizes range from 1 to 5 mm (0.04 to 0.2 in.) with 3% to 4% chalcocite as the chief ore mineral. The rock contains numerous open and healed fractures. EXPERIMENTAL DESIGN AND SHOCK WAVE THEORY Figure 1 shows the inclined plate plane-wave generator used to induce a uniaxial shock wave in a cylindrical specimen of copper porphyry (20.3 cm (8 in.) in diameter and 10.2 cm (4 in.) in height). The "mousetrap" assembly, described by Benedick (1972), uses a triangular line- wave generator and Detasheet (1) of varying thicknesses as the main explosive charge in direct contact with a 30.5 cm (12 in.) square plate of 2024 aluminum alloy.
- Geology > Mineral > Native Element Mineral > Copper (1.00)
- Geology > Mineral > Silicate > Tectosilicate > Quartz (0.67)
- Energy > Oil & Gas > Upstream (1.00)
- Materials > Metals & Mining (0.89)
ABSTRACT INTRODUCTION Attempts to analyze the stability of slopes, foundations and underground openings in discontinuously jointed rock have generally assumed full joint continuity or ignored the role of stress concentration at the tips of discontinuous joints. The most common approach in rock slopes is to consider a failure plane consisting of coplanar, but discontinuous joints separated by intact rock. The intact rock is assumed to fail in shear as a Mohr- Coulomb material and to contribute a strength component proportional to the fraction of the failure plane that is intact rock (Jennings, 1970). An alternative approach by Stimpson (1978) applies Lajtai's (1969) theory for shear strength of discontinuously jointed rock masses to slopes. Lajtai allows tensile failure in the intact rock of a composite surface. This theory yields lower shear strength than that of Jennings (1970) except in the presence of high normal compression, however it also fails to account for the stress concentration at the tip of a discontinuous joint. For a given slope geometry, stresses increase with slope height, but in both of these models rock mass strength is independent of size. A fracture mechanics analysis shows that particularly for small slopes, rock mass strength decreases with size. A continuous stepped failure surface produced by propagation of preexisting joints may develop at lower shear stress than predicted by the Jennings (1970) or Lajtai (1969) models. This fracture mechanics model applied to slopes in competent rock is more realistic in its assumed failure mechanism, and for large slopes more conservative than the Jennings (1970) or Lajtai models (1969). To understand the behavior of slopes, both tensile and shear strength must be considered, because both may resist movement in discontinuously jointed rock. Tensile strength criteria govern the zone of tensile stress and separation at the head of the slope, while shear strength provides resistance over the balance of the slope. TENSILE STRENGTH Rock masses are typically assumed to have no tensile strength, but significant strength is possible where joints are discontinuous. For a single isolated crack the uniaxial tensile stress applied to the rock mass for incipient crack propagation is (Rooke and Cartwright, 1976): (mathematical equation) (available in full paper) where KIC is critical stress intensity and a is half the crack length. If the conservative assumption is made that tensile failure results immediately from crack propagation, s t defines the tensile strength of the material. For strong rock Kic varies over a relatively small range and a value of 1.0 MNms assumed for purposes of calculation. For long-term loading this value may be reduced to account for stress corrosion (Anderson and Grew, 19771Atkiusou, 1982). For widely spaced joints 10 cm long, the tensile strength of the rock mass is 2.5 MPa (370 psi), a substantial fraction of the strength of strong intact rock. Even for joints 10 m long the strength is 0.25 Mpa (37 psi). Based on Westergaard (1939) and Rooke and Cartwright (1976), the strength s t for a rock mass with tension applied perpendicular to an array of equally spaced, identical coplanar cracks is: (mathematical equation) (available in full paper)
- North America > United States (0.47)
- Africa (0.28)
ABSTRACT ABSTRACT The effectiveness of coalface powered support depends upon its structural nature and the manner of application of the hydraulic system. A critical and largely unresolved feature of the system is the optimum setting pressure level at which a support should operate in its geological environment. This paper provides an argument for a generalized approach to the issue of setting pressure level quoting case study data for a high output British longwall installation. INTRODUCTION The uniqueness of each mining location brings into question the validity of a generalized approach to mine design. Despite the wide range of currently available hydraulic supports, it may be argued that strata control at the coalface has successfully been standardized. Design criteria have been produced such that a wide range of geological and mining environments experience controlled strata yield. The present generation of coalface support forms part of an evolutionary process commenced between 1930-1950, and which has since developed rapidly as a result of technological advances in all the major branches of engineering. The general level of efficiency of mining systems is high with regard to manpower, machine utilization, and mine layout. A persistent number of low efficiency powered support installations operate in conditions of large roof and floor debris layers, inadequate support and roof bed fracturing, and slow face advance - a situation that can be accurately described as a 'vicious' cycle. FIELD INVESTIGATIONS South Wales Coalfield Research at Cardiff between 1972-1981 at numerous locations in the South Wales Coalfield, confirmed a characteristic loading and convergence pattern for the cycle of extractive operations in which support advance/reset appeared as the major influence upon roof- floor convergence. The introduction in 1981 of a transducerised data logging system for data acquisition at coalface investigations of support and strata behavior (Isaac et al, 1982), enabled monitoring of various parameters to be virtually continuous in nature. The pattern of loading and convergence was observed to be similar as before, except that a pressure decay effect was observed in the early stages of support load development. This was explained as the effect of roof and floor debris compaction at support reset. For the relatively low setting levels of British installations at between 10-15 MN/m, this period of pressure decay was accompanied by a high rate of convergence with only a small amount of leg closure. This behavior suggests uncontrolled load development at the critical period of support advance/reset in the extractive cycle. The implication of this condition is that roof bed strata would suffer induced fracturing, a greater amount of roof-floor convergence would occur, and that powered support performance as expressed by the Pressure/Convergence criterion (Smart and Isaac, 1982) would be less efficient. Nottinghamshire Coalfield Investigations at Bilsthorpe Colliery, North Notts. Area, National Coal Board, in June 1982 and April 1983 (Isaac and Bradbury, 1983) provided the opportunity of assessing Gullick Dobson chock-shield supports. The initial study of 4 leg/260 tonne supports was conducted at 23's Face in the Park gate Seam (1.8m thick, shallow dipping), the second study being performed on 4 leg/280 tonne supports fitted with 'positive set' valves, at 15's Face in the same seam.
- Europe > United Kingdom > Wales (0.95)
- Europe > United Kingdom > England > Nottinghamshire (0.24)
Comprehensive Observation And Research On The Mechanism Of Water-Irruptions From The Floor
Zigang, Jin (Special Coal Mining Research Section) | Baiying, Li (Special Coal Mining Research Section) | Zhenpeng, Sun (Coal Scientific Research Institute, Fengfeng Bureau of Mines) | Zhenan, Wang (Coal Mining Engineering Department, Shandong Mining College)
ABSTRACT ABSTRACT In China there are more than 30 coal fields menaced by karst-confined water. In the north and east of China there are many collieries which mine coal seams of the Carboniferous and Permian Systems, which are above the Ordovician limestone (general thickness is 500-600 meters). Ordovician limestone is full of karst fissures, contains a lot of water under high pressure (15-65 kg/cm) near the commercial coal seams (20-80 meters), the water easily infiltrates the thin limestone at the point of the structure crevice, so water-irruptions from the floor are the main source of water-related disasters in these coal fields. According to the statistics there were more than 300 accidents of all kinds of water-irruptions from the floor in these fields since the beginning of their development. It is of great significance in developing coal production in the eastern part of China to solve the problem of safe mining of the seams menaced by the karst-confined water of the Ordovician limestone. The research on this subject consists of three parts: (1) investigating the geological structure and hydrogeological conditions of each mine of the mining area; (2) researching the mechanism of water-irruptions from the floor and proposing reliable methods to predict and monitor; (3) improving the mining system and proposing a mining plan with safety precautions against water-caused accidents. The chief task is to research the mechanisms involved and to discover the causes and conditions of the phenomenon. The authors have obtained some findings on this subject in the last five years, making use of the comprehensive observation and research system, which will be discussed under three headings. CLASSIFICATIONS OF WATER-IRRUPTIONS FROM THE FLOOR Each water-irruption from the floor has a different cause. In order to determine the causes and conditions of floor water-irruptions we must first classify them as follows. First, according to the position of the irruption, it can be classified into two types: (1) at the excavation opening; (2) at the working face. According to field statistics, these two kinds of water-irruptions account for 50% each of the total. Most of the irruptions at the excavation opening are under conditions of meeting with permeable faults or the structure crevice zone connecting with the karst-confined water, and the probability of an irruption is greater when the opening is near the fault, or while excavating at the fault. To prevent water-irruptions at the point of excavation, the water-discovery excavating method must be used with caution. The mechanism of water-irruptions at the mining face is very complex and will be discussed separately. Second, according to the manner of the floor water-irruption, it can be classified into three types: (1) sudden bursting; (2) slow bursting; and (3) delay bursting. Sudden bursting usually occurs near the excavating and mining faces. As soon as the water-irruption occurs, the water inflow reaches the maximum value (peak) in a short time, the irruption is very strong at high speed, the immediate floor of the seam will lift, and the water will carry rock fragments. Slow bursting also occurs near the excavating and mining faces, but water inflow increases gradually and reaches the maximum value in several hours or days or even several months.
ABSTRACT ABSTRACT In support of the DNA Deep Basing Program, SRI International has been performing scale model experiments on protective structures since 1973. The SRI program includes field and laboratory tests of mini- structure models with surrounding natural or simulated geology. This paper contains a synopsis of some of the results of this research. INTRODUCTION With the suspension of multiple-aim-point basing of the Peacekeeper missile and the growing need for endurance of stategic forces, deep basing is being explored as an alternative basing mode. Also, deep based underground structures have been under continuing investigation as a method for hardening command and communications centers. These investigations have resulted in the need for test data on protective structure response to nuclear attack loads. Testing of protective structures in underground nuclear tests (UGT) is very expensive and time consuming, and experimental results are not easily extrapolated from one geology to another. Perhaps more important, field measurements of response in the rock field around the structure are extremely difficult to make. Even with great effort, observations are limited to a few isolated points accessed by means of bore holes. There can be no direct observations of rock flow and fracture fields, or displacements along joints and bedding planes. Because the rock is the essential load-bearing member, these limitations of field testing have severely hampered protective structure design. In addition, structure response at load levels higher than those attainable in field tests is often of interest. The use of scale models to test the response of protective structures is an approach which can overcome these difficulties. FIELD TESTING In the MIGHTY EPIC event [Refs. 1,2], small-scale tunnel structures (nominal 150-mm-diameter tunnels) were fielded directly in the native tuff. The tunnels were reinforced with steel liners, aluminum liners, and steel liners with crushable urethane foam backpacking. They modeled larger prototype structures fielded in the same event. Quantitative comparisons of tunnel closures showed the same percent closures in structures of the same strength for both the mini-structures and the larger scale structures fielded in the same drifts. The back-backed structures at both sizes performed very well and demonstrated that reinforcement designs that allow a few percent tunnel closure, and hence yielding in the rock mass, can carry substantially more load than tunnels that must remain elastic (Figure 1). Fig. 1. Comparison of build-up structure SX-11 that survived and composite integral structure that failed at the same load. (available in full paper) Also, the extensive closure data from the mini-structures showed a smooth increase in closure with increase in load-to-strength ratio for data taken from the three drifts. Thus, the tuff behaved mainly as a uniform homogeneous material throughout the test bed, and exhibited the same strength properties at both small and large scale. Mini-structure models of tunnels in modeled jointed rock masses were fielded in the DIABLO HAWK underground nuclear test [Ref. 3]. The simulated jointed rock masses, as well as the structures, were recovered after the test and sectioned to examine the complete response of both rock and structure.
- North America > United States > Kansas > Butler County (0.24)
- Asia > Middle East > Israel > Mediterranean Sea (0.24)
- North America > United States > California > San Mateo County > Menlo Park (0.17)
ABSTRACT INTRODUCTION Purpose of Wall System There were 2379.88 m (meters) of tied back wall along Interstate 77 in the Capitol Complex area in Charleston, West Virginia. The main purposes of this wall system was for aesthetics, reduction of excavation, saving the forested slopes, and control of sloughage of the back slopes onto the Interstate. Failure At 1 PM on Thanksgiving Day, 1982, 3 wall panels fell from the insitu position onto the interstate in a matter of seconds. See Fig. 1. The total length of wall that fell was 29.25m. The height of the three panels varied from 10.06m to 10.36m and each individual panel had a length along the highway of 9.75m. These were abutting panels that were approximately in the middle of a section of wall 165.81m long. The panels separated from the remainder of the wall at 2 expansion joints that occur at the ends of every 3rd panel. The reinforced portland cement concrete panels were designed on a 76.2mm (millimeter) batter. See Fig. 2 for Wall Dimensions. The footer for these panels was 1m wide, was keyed into sandstone bedrock 0.30m and was approximately 1.52m below the portland cement concrete pavement and shoulder. Thirteen of the 17 panels in this section of wall were tied to the sandstone bedrock behind the wall by either 25.4mm or 34.93mm hollow rock bolts. Two panels on each end of the wall were cantilever type walls. The panels that failed had 25.4mm hollow rock bolts. A portion of each panel adjacent to both sides of the failure was pulled out from the rock face and appeared to be in imminent danger of also failing. All of the 21 rock bolts in the panels that fell were separated either at the backface of the wall, in the extension from the wall to the rock face or at the rock face. None of the rock bolts had been pulled out of the bond length in the sandstone bedrock. After considering the condition of the wall and rock bolts, it was decided that the remainder of this section of wall (14 remaining panels) should be taken down. Attempts to take the remaining wall down, even the 2 panels adjacent to the 3 that failed, were unsuccessful until almost all of the existing rock bolts that were still intact were cut and large construction equipment used to push these panels over. The 3 north bound lanes of 1-77 were closed during the 3 days of around the clock removal of the wall. No personal injuries occurred as a result of the failure, although one automobile was narrowly missed by the failing wall. Method of Investigation The technical investigation was assigned to the Soil and Rock Mechanics Section of the Materials Control, Soil and Testing Division of the Department of Highways assisted by D. R. Piteau and Associates. Dr. Carl D. Lundin of Materials Application Inc was employed to evaluate the corrosion of the bolts. The investigation was to determine the cause of the failure and if it was attributable to external causes. It was also to be determined immediately if the failure was the result of bedrock or overburden movement.
- Construction & Engineering (1.00)
- Transportation > Ground > Road (0.88)
- Automobiles & Trucks (0.88)
- Well Drilling (0.90)
- Facilities Design, Construction and Operation > Pipelines, Flowlines and Risers > Materials and corrosion (0.68)
- Reservoir Description and Dynamics > Reservoir Characterization (0.50)
- Well Completion > Well Integrity > Subsurface corrosion (tubing, casing, completion equipment, conductor) (0.46)
ABSTRACT ABSTRACT Examples are presented in which Time Domain Reflectometry (TDR) was employed to locate deformation in rock masses induced by mining. The first example involved monitoring the propogation of overburden fractures above a longwall coal panel and the second example involved the use of TDR to monitor rock mass deformation in the vicinity of a strip mine highwall. Rock mass deformation, strata separation and large block movement generated strains and failure along coaxial cables which were installed in drill holes and the TDR reflections generated by these strains and failures were monitored. Analysis of field data and a preliminary laboratory test indicate that it may be possible not only to locate rock mass movements but also to quantify these movements with TDR data. INTRODUCTION Time Domain Reflectometry (TDR) is an electrical pulse testing technique originally developed to locate breaks in power transmission cables. Recently, this technique has been adapted for use in monitoring the movement of rock masses during mining. Coaxial cables have been installed in drill holes or along mine entries such that rock mass movement will damage the cables. When electrical pulses are transmitted along the cables it is possible to locate rock mass movements by virtue of the pulse reflections created by cable breaks. TDR OPERATING PRINCIPLE AND INSTALLATION A TDR installation consists of a TDR cable tester and the cable to be tested, Ultra-fast rise time voltage pulses are sent down the cable from the tester, Cable defects such as crimps, short circuits or breaks reflect these pulses back to the cable tester, The returned voltage, sometimes called the reflected voltage, is superimposed on the advancing initial step and will appear as a step- up or step-down transition on the cable tester display as shown in Figure 1, Inductive faults or faults of higher resistance than that of the cable (such as a frayed or cut cable) cause reflections in phase with the initial step resulting in a step-up transition, Capacitive faults or faults of lower resistance than that of the cable such as a short circuit or crimping will cause out-of-phase reflections resulting in step-down transitions, Assuming a constant pulse propagation velocity, the distance to the cable fault is proportional to the elapsed time between initiation of the voltage step and arrival of the returned voltage. The cable tester measures this time in units of distance. Consequently, it is possible to determine the location of a cable fault and the nature of the fault by inspection of the TDR signal that is recorded. The capability of distinguishing cable breaks from cable crimps provides a convenient means for improving the accuracy with which cable breaks can he located. This accuracy is on the order of 2Z of the distance from the tester to a cable break so that a break at a distance of 30 m (100 ft) can only he located to within 0.6 m (2 ft) of its actual position. If a cable is crimped at selected intervals however, a set of reference marks is recorded each time the cable is tested. The location of cable breaks or any other cable defects which result from rock mass movement can then he determined with respect to these reference marks.