In an effort to gain a better understanding of shale Joint behavior, normal-only and shear creep tests were performed on artificially prepared shale Joints. Pro- and post-test surfaces were examined for changes in morphology. This paper discusses both of these programs and the results derived from them.
Changes in stress may occur as a result of construction, excavation or the release of residual ground stresses. When the stress regime is altered there is a redistribution of stress throughout the rock mass. The process of stress equilibration includes both 'instantaneous' and time-dependent components. In some instances the cumulative time-dependent strains may exceed the instantaneous elastic response of the rock, by several orders of magnitude. The possibility of this occurrence necessitates that creep movements be considered in the design of surface and underground structures constructed in or on rate-sensitive materials. The strains associated with the readjustment of stress may be manifested as creep along the Joints and/or creep of the intact material. Both types of responses have been observed for shaly materials (Quigley et al, 1978; Lee and Klym, 1978; Bowden, 1982). The total response of a rock mass depends on the relative contribution from each source. Joint creep, is the time-dependent relative displacement of opposing Joint surfaces, caused by a stress smaller than the shear strength of the joint. Our concern for the creep behavior of joints can be justified on the basis of their pervasiveness and the fact that the shear strength of a joint is generally less than that of intact rock. As such, joint creep can account for a significant proportion of the total deformation of a rock mass. Numerical methods are available which can model jointed rock masses, including the time-dependent behavior of both the intact material and joints (Curran and Crawford, 1980; Crawford and Curran, 1983). However, in order for these models to be realistic, we must know the stress-strain-time characteristics of the discontinuities. Furthermore, individual discontinuities which may decisively influence the properties and behavior of the rock mass, must be correctly represented geometrically and be accorded strength and deformation characteristics which reflect their behavior in the field. The accuracy of the stresses and displacements that a model yields is proportional to the quality of the numerical relationships which we use to approximate actual joint behavior. Our ability to correctly and consistently determine these relationships depends on the number and types of factors which influence a joint's behavior and our understanding of them. Shale joints are particularly difficult to model because their frictional strength is, among other things, time-dependent, and we lack a basic understanding of shale joint phenomena. This paper discusses the response of artificially prepared shale joints to normal only, and combined normal and shear loading. The implications of the observed behavior on design criteria are also examined. A large scale, high load capacity single surface bi-directional shear machine (sample size 200 x 300 mm) was designed and constructed at the University of Toronto for the purpose of studying the creep behavior of shale joints. Plate 1 shows the assembled machine and support equipment prior to testing.
A series of cross-hole acoustic measurements have been performed in a columnar-Jointed basaltic rock mass around an underground opening mined by the drill-and-blast method. The purposes of the test program were: to evaluate the rock mass characteristics around the opening; to determine the zone of blast damage; and to evaluate seismic methods for anomaly detection ahead of mining in this type of rock. The cross-hole measurements were made between four 76-mm diameter horizontal boreholes diamond-drilled 12 m into a wall of the underground opening. Repetitive pulses of compressional (P) and shear (S) waves of frequencies in the range 1 kHz-100 kHz were propagated from a transmitter sonde through the rock mass to a receiver sonde, both of which were clamped hydraulically to the borehole wall. After amplification the received P- and S-wave signals were digitized at the surface by a digital oscilloscope and stored on floppy discs. The results indicate considerable reductions in P- and S-wave velocities at distances less than 2 m from the face. Clearly these low values are associated with blast damage. Beyond 2 m, the velocities in a vertical direction indicate almost constant values. The velocities in the horizontal direction beyond 2 m appeared erratic, but showed a general tendency to increase as a function of distance from the face. Their maximum values remained, however, still lower in value than the corresponding velocities in the vertical direction. Near the face, the differences in velocities were considerably greater: with horizontal velocities considerably lower in value than those vertical. Results of the spectral analyses of the received signals indicated that Q-values were strongly influenced by the vertically- oriented sets of Joints. Only for waves travelling in the vertical direction was the effect observed of the blast-damaged zone immediately around the opening. The acoustic data obtained are clearly indicative of an anisotropic jointed rock mass, with a greater intensity of Jointing for travel paths in the horizontal than the vertical direction. The vertical, columnar Joints are probably less tightly closed than those oriented in the horizontal plane.
The in situ assessment of geomechanical characteristics of rock masses is an essential prerequisite to the design and analysis of major structures, both on the surface and underground. A particular in situ investigative technique that has shown considerable promise for this purpose is the cross-hole acoustic method. The velocities of compressional and shear waves and their attenuation in rocks containing fissures, fractures and joints have been observed to be influenced strongly by the state of stress, changes in temperature, and degree of water saturation in the rock mass. Nut and Sizenons (1969) have shown that the elastic properties of rocks are controlled mainly by the properties of microcracks present at low stresses, and that the application of uniaxial stress caused elastic-wave anisotropy, with a higher compressional-wave velocity in the direction of the applied stress. King (1969) and Anderson et al (1974) found that a preferred orientation of open cracks had a marked effect on seismic velocities, with the major reduction in velocity observed perpendicular to the plane of fractures.
A rapid in situ stress measurement technique was developed for exposed underground surfaces. The method applies radial stress in small diameter (38mm) boreholes, initiating fracture propagation which is monitored by measuring ultrasonic pulse velocity. Theoretical relations for stress concentrations near boreholes provide independent expressions for unknown principal field stresses at incipient fracture; fracture orientation defines principal stress orientation. Accuracy and precision were evaluated using blocks of brittle material subjected to known stresses. Results were comparable to those obtained by other in situ stress measurement techniques.
The U.S. National Committee for Rock Mechanics reported a need to make existing test methods more cost effective. In situ stress data is rarely used because of high cost and questionable accuracy. Some in situ stress measurement methods require costly non-recoverable components; others demand tedious slot cutting or overcoring. In some materials, usable results come from only 20% of the tests performed (Van Heerdon and Grant, 1967). This paper presents preliminary work toward a rapid, accurate technique for in situ stress evaluation at a free surface. The method may also serve in deep hole applications.
In situ stress by induced fracture involves assessment of pressures required for incipient crack formation on the periphery of a borehole. Crack initiation during borehole pressurization is controlled by the material's tensile strength and in situ field stresses at the borehole boundary. However, cracks reopened upon repressurization are no longer dependent on tensile strength, so stress concentration theory taken with known borehole pressures as cracks are caused to reopen provides information about in situ field stress. A two stage process providing information sufficient to calculate both principal stresses in a plane perpendicular to a borehole axis is proposed. Cracks are first induced from a single borehole, where theory predicts fracture should propagate in a direction normal to the minor principal stress (Hubbert and Willis, 1957). Next, simultaneous pressurization of a pair of bore- holes placed close to each other and aligned parallel to the minor principal stress is performed. This should result in a fracture normal to the major principal stress. Theories for stress concentration near single and double holes provide two relations in terms of the two unknown principal stresses.
Jeffrey (Timoshenko and Goodier, 1967) showed that the tangential component of a uniform radial pressure, Pi, acting on the inside boundary of a cylindrical hole is: (mathematical equation) (available in full paper)
The negative sense of Eq. 1 indicates tension. The tangential stress component as provided by Kirsh (Timoshenko and Goodier, 1967) for field stress concentrations at the wall of a single hole is: (mathematical equation) (available in full paper)
where: s1 = major principal stress, s2 = minor principal stress and ¿ = angle measured from the direction of major principal stress. For unequal compressive principal stresses, Eq. 2 shows minimum compressive stress occurring at ¿ = 0 and p as: (mathematical equation) (available in full paper)
The G-Tunnel heated block experiment is being conducted on the Nevada Test Site (NTS) as part of the Nevada Nuclear Waste Storage Investigations project (NNWSI). The purpose of the ambient temperature testing phase is to evaluate rock-mass mechanical properties of a block (~8 m3) under biaxial stress changes up to 7.5 MPa above an initialization in situ value of 3.1 MPa. Results indicate that the modulus of deformation ranges from 9.7 to 17.0 GPa and Poisson's ratio ranges from 0.21 to 0.33. In general, the higher values of the modulus and Poisson's ratio were influenced by fracture propagations parallel to the compressive stress field. Other measurements indicated that cross-hole compression (p) wave velocities and single fracture permeability values were relatively insensitive to stress changes above the in situ value.
Volcanic tuffs are being considered by the Department of Energy (DOE) for the possible disposal of commercial high level radioactive wastes. The Nevada Nuclear Waste Storage Investigations (NNWSI) project was established in 1977 to evaluate such disposal in geologic formations on or adjacent to the Nevada Test Site (NTS). Sandia National Laboratories (SNL), as one of the participants of the NNWSI project, is responsible for the rock mechanics program to support the design and evaluation of a repository in tuff. The data from this heated block experiment provide key input for definition of the rock-mass mechanical properties to be used in the conceptual design. The rock mechanics field program has begun in G-Tunnel, where tuffs similar to those at the potential candidate site are found; later experiments are planned as part of the exploratory shaft investigations at Yucca Mountain, the potential site. Science Applications, Inc. (SAI) has been under contract to SNL to provide assistance in the design, installation, operation, and evaluation of this experiment. They have been responsible for the preparation and installation of the instrumentation and the control systems as part of this effort. Related details will be presented separately from this paper. The purpose of this paper is to present the results from rock-mass mechanical measurements taken prior to the application of heat to the block. The heated block is a multi-faceted experiment designed to evaluate rock-mass thermal (thermal conductivity), mechanical (modulus of deformation, Poisson's ratio, joint shear and normal properties), thermomechanical (coefficient of thermal expansion), and hydrological (fracture permeability, moisture content, pore pressure) responses on a relatively large scale (~8 m3) for conditions where stress and temperature boundary conditions can be reasonably well controlled. The heated block is located in the floor of an alcove driven from the Rock Mechanics Drift in the G-Tunnel Under- ground Facility. The alcove is in welded tuff and was mined using standard drill and blast techniques. The block was formed by cutting four 3-m-deep orthogonal slots in the floor of the alcove to form the periphery of a prism of jointed-welded tuff 2 m square. The slots were cut with an overlapping line drilling method using percussion drilling techniques. Flatjacks were grouted into the slots in configurations shown in Figure 1.
Computer assisted finite element model analysis, originally developed to study stress distributions in aircraft, has been modified into a versatile mining program, that can be used to assist the engineer with predictions of mine subsidence. International Exploration, Inc. has developed an in-house finite element model program which incorporates a multitude of factors influencing mine subsidence such as mine design, physical rock properties, structure, faults, joints and topography. Finite element analysis of a mine model can mathematically determine the magnitude and direction of the principal stresses within each of the elements forming the model. By comparing the calculated in-situ stresses to the actual rock strengths, high risk regions within a mine can be identified. These potentially unstable regions of the mine can be allowed to fail within the model and the resultant stress redistribution and element movement calculated. Thus, the effects of the mine collapse at the surface can be predicted for the purpose of determining the potential damage, if any, that could occur to surface structures from deep mine failure. Stabilization materials can also be mathematically inserted into the model to design and analyze the effectiveness of subsidence prevention programs. Two examples are presented of the finite element program being used' to predict mine subsidence at substantially different geologic and mining provinces in Pennsylvania.
Abandoned subsurface coal mines in Pennsylvania have posed a substantial problem to hundreds of property owners situated above them. The mines may collapse, with time, from various causes such as weathering, pillar robbery or mine flooding. The resulting subsurface collapse may cause significant surface movement of a magnitude and direction capable of damaging surface structures. As a result, mine stabilization projects are faced with determining which areas within hundreds of square miles warrant special consideration. Through the use of computer assisted finite element modeling, specific high risk areas can be delineated and recommendations suggested for developing strategies to prevent future damage to surface structures. Using finite element modeling, the mining and geologic conditions of a specific study site are mathematically recreated to calculate the stresses present within mine pillars and rooms. Over stressed areas within the model can then be caused to fail and the resulting magnitude and direction of surface movement and stress then determined. The mathematical model can be modified to consider the effects of pillar decay and robbery as well as introduce faults, secondary mining and stabilization materials. The ability of the program to calculate the stress conditions within a mine under the complex situations commonly found in Pennsylvania greatly exceeds the capabilities of empirical approaches.
FINITE ELEMENT MODELING
The finite element method is a general mathematical technique of structural analysis in which the body to be studied is subdivided into individual structural elements, in this case, triangles (Zienkiewicz, 1977) (Figure 1). The triangles are interconnected at .their corners or nodes. When forces are applied to the body, the nodes displace, and the triangles experience strain. The amount of displacement of each triangle depends on the level of the forces applied to the triangles and the material properties of each triangle.
Rock-mass fracturing was measured on oriented drill core and along scanlines in three mines extracting mineralized porphyry deposits. Fracture families identified in the core were consistent with those found along scanlines, but fracture frequency in the core was twice that measured along scanlines. The excess fractures, which occur within 45 ° of a right angle to the core, apparently are produce during drilling, which implies that fracture spacing in drill core is an ambiguous quantity dependent on the drilling procedures, on the orientation and cohesion of the planes of weakness, and on the actual fractures present in the rock mass.
One objective in logging and testing drill cores is to infer the structural characteristics of the rock mass. The structural stability of an excavation in rock presumably depends in part on the geometry of fracturing or jointing in the rock mass--the number of distinct fracture families, their orientations, the spacing between fractures, and the extent or size of the fractures. Within any fracture family the orientation, spacing, and extent are never fixed values; rather they vary over substantial ranges, just as do the strength and stiffness parameters. The variation is usually describable in terms of a frequency distribution, most commonly the negative exponential or the lognormal. Determining the geometry of rock-mass fracturing from drill core is not a simple task. Unless the core is oriented on recovering it from the drill hole, the fracture families cannot be established, and hence the fracture spacing can be determined only between successive fractures along the core, without regard for fracture family. Fracture extent cannot practically be determined from core that is only 30 to 50 mm in diameter. Nevertheless, many instances arise in which exploratory drill cores, ordinarily unoriented, are available for study. Fracturing measurements from oriented cores and scanline mapping, summarized herein, are used to explain some of the fracture distributions that may be expected in unoriented cores.
DATA ACQUISITION AND RESULTS
Jointing geometries were determined for three ore bodies being mined by undercut-cave methods (Panek and Melvin, 1984), part of an investigation aimed at calculating the cavability of a large mineralized rock mass as a function of the jointing and of the strength and stiffness parameters (Panek, 1981). The San Manuel is a typical porphyry-copper ore body, strong but well fractured; the Lakeshore mine copper-oxide deposit is highly altered and friable porphyry and metasediments; the Henderson mine molybdenite deposit is a competent, tight, granite rock mass. The three deposits provide a wide range of strength characteristics; representative unconfined compressive strengths are 90, 30, and 107 MPa, respectively. At each mine the rock mass was sampled by core drilling three or four NX-size holes in different directions at each of one or two sites, producing 180 to 300 m of core by a different pattern of holes at each mine. A split, double-tube core barrel was used, and hence core recovery was virtually 100%. The core was oriented by making an impression of the core stump remaining in the hole after the removal of each length of core and matching the impression to either side of the break surface.
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:
Cutter roof failure is a specific type of ground control problem which frequently results in massive roof failure. It is a common occurrence in coal mines of the Northern Appalachian Coal Basin, causing delays in production and posing a safety hazard to mine personnel. The Bureau of Mines is conducting research on the causes of cutter roof failure to gain a basis from which to prevent its occurrence and to support such roof when failure does occur. Research conducted in a coal mine of central Pennsylvania has revealed a correlation between the occurrence of clastic dikes and formation of cutter roof failure. In-mine mapping of ground conditions showed an increase in roof failure in areas of high frequencies of clastic dikes. Rock pressure monitoring around clastic dikes registered the greatest amount of roof loading near the intersection of dikes with the rib. Load cells measuring horizontal pressure changes in the roof indicated that the greatest pressure changes were occurring perpendicular to entry headings when clastic dikes were present. Analysis of rock pressure monitoring shows that the roof behaved as two cantilever beams when severed by a clastic dike. Additional roof supports such as trusses and cribbing were found to effectively support the roof in areas of clastic dikes and prevent cutter roof failure. However, these methods were only successful when employed shortly after mining.
Occurrences of cutter roof failure during development mining have been observed in increasing numbers. In response to this, the Bureau of Mines is investigating the causes of this type of roof failure, which can deteriorate to a major roof fall if additional support is not applied. The uniqueness of this problem has limited the techniques available for its treatment. The approach of trial and error treatment has proven unsatisfactory in many cases, as the causes are often different from mine to mine. Cutter roof failure initially begins as a fracture along one or both roof-rib lines of an entry and propagates nearly vertically into the roof (fig. 1). When the fracture breaks to a height above the anchor horizon, or along a weak bedding plane, massive roof failure may occur. Figure 2 is an example of severe cutter failure and figure 3 illustrates the end result of such cutter development. Previous research by the Bureau has shown a correlation between cutter roof failure and high horizontal stress fields. Aggson (1979) and Kripakov (1982) conducted studies addressing the particular stress state that would initiate cutter roof using finite element analysis. Research conducted by Thomas (1950) approached cutter development from a purely practical perspective, without any instrumentation. His work is believed to be the earliest attempt to understand cutter development. The influences of the direction of mining and occurrence of clastic dikes on cutter roof failure formation have been observed through detailed mapping by Iannacchione, et.al., (1984). The occurrence of cutter roof has had a major effect on entry stability at the Greenwich Collieries North Mine of Indiana County, Pennsylvania. The investigation conducted there was comprised of two basic phases which contribute to an understanding of the pres- sure dynamics surrounding cutter development and the influence of geologic anomalies.
A combined numerical approach is proposed for the analysis of rock-liner interaction in tunnels under the effects of severe dynamic loads. The method is based on employing finite element formulations which describe the linear behavior and incorporating it into a finite difference code which can describe the behavior of a jointed rock medium. This method can be applied for the analysis of tunnel systems in rook, and for studying the requirements for adequate support of the tunnel.
Tunnel performance under dynamic loading, which may arise from earthquake or blast induced motions, is of significant interest to engineers who design tunnels, and to researchers who study tunnel behavior. In general there are two principal types of tunnels in rock: