ABSTRACT: Numerical simulations of hydraulic micro-fractures are used to provide insight into techniques commonly used for stress measurement at depth. These simulations encompass initiation and propagation of hydraulic fractures from a borehole in poroelastic rock. It is shown that poroelastic effects may have a significant influence on the determination of the principal stresses in permeable rock. Specifically, it is found that (1) the breakdown pressure is not necessarily associated with the fracture initiation at a borehole, (2) the time of fracture closure can be identified from borehole pressure logs and (3) poroelastic effects can cause the borehole pressure at the time of fracture closure to markedly exceed the minimum in-situ stress
There has been a considerable effort directed towards practical development of techniques for in-situ stress measurement using hydraulic fracturing. However, there have been relatively few detailed simulations of these processes [2 ,3]. A series of simulations of hydraulic fracturing tests is presented. Here it is assumed that the rock mass is a poroelastic material. This assumption requires that the following phenomena be treated as fully coupled: (1) the fluid flow in the fracture,( 2) the deformation of the rock mass, and (3) the fluid flow in the rock. Treatment of the rock in this manner is a unique feature of the simulations presented herein.
A detailed description of the numerical procedure used in these simulations can be found in Ref. 4. Its salient characteristics are as follows: (1) the finite element method is used to approximate the fully coupled poroelastic solution for deformation and fluid flow in the rock mass, (2) a finite-difference approximation is used to model the fluid flow in the fractures, and (3) a generalized Dugdale-Barenblatt fracture model is incorporated as a natural product of the solution procedure. This fracture model allows the crack length to increase or decrease through the course of the simulation so that events related to shut-in and fracture closure can be investigated.
The results of these simulations are presented in the next section. In the third section, their significance is described and discussed with reference to published experimental and field data.
2. A SIMULATION OF FRACTURE INITIATION, PROPAGATION, AND CLOSURE IN POROELASTIC ROCK.
Hydraulic fracturing for the purpose of stress measurement involves the following: (1) pressurization of a borehole until a fracture is initiated; (2) controlled flow of fluid into the borehole to propagate the fracture a short distance; and (3) shut-in or cessation of the flow into the borehole, followed by monitoring of the pressure decline.
Figure1 . The finite element mesh used in the simulation of a series of hydraulically driven fractures in poroelastic rock. The fracture is constrained to propagate along the x-axis. (available in full paper)
The crack-mouth-pressure (CMP) versus time curve is then used to deduce the in-situ stress state. In this section, results are presented for detailed simulations of initiation and propagation of a hydraulic fracture from a borehole in a poroelastic material.
ABSTRACT: Dynamic rock anchors are interior fixtures developed by the author which promise to be a revolutionary development in the field of ground control. The fixtures are designed to be placed in the ground with modern roof bolting equipment, such as the type used in the American coal mining industry. They are composed of a plastic formable anchor tube which serves to fasten a buttress deformed rod which is thrust and turned into the plastic and in turn clamps the roof plate to the rock surface. Plastic tubes 10 to 16 inches in length provide sufficient anchorage to exceed the yield strength of steel rods 5/8, 3/4 or 7/8 inches in diameter.
This paper presents field testing and early work done on the device over the last seven years. Results of pull tests, creep tests, laboratory testing in concrete, and field testing under natural rock conditions demonstrate the uniqueness and effectiveness of this new support system.
A second development which uses the dynamic rock anchor in conjunction with a resin cartridge to provide a stiffer anchorage system is also discussed. These products will be manufactured and marketed by Ingersoll-Rand Co. under the trademark of DYNA-ROK and DYNA-ROK PLUS anchors.
DYNA-ROK anchors have resulted from an extensive research effort. From 1981 to 1984 the author did what he terms "dirty" research to separate various anchor devices and systems. This resulted in the selection of the plastic deformable anchor as being the best candidate for success.
Once testing proved that the anchor was feasible, tests were conducted at the Spokane Mining Research Lab. This resulted in a refinement of effort as electronic readouts could be made of the torque, anchorages achievable, and plate loads.
In 1986 an agreement was finalized with Ingersoll-Rand Co. which did further refinement, research and development in their labs. Ingersoll-Rand is the licensed manufacturer of the product. DYNA-ROK and the DYNA-ROK PLUS are registered trademarks of Ingersoll-Rand for these products.
To install DYNA-ROK fixtures, modern mechanized roof bolting equipment typical to the American coal mining industry is used. The concept is simple. The hole is drilled, the tube inserted and the fixture thrust and torqued to tension the rod and seat the roof plate. The rock surrounding the anchor is strengthened by active spiral thrust vectors from the buttressed rod through the plastic and into the rock.
It is a long point anchor system with the anchor extending over sufficient length so that loads against the rock are not great. The plastic anchor tube eliminates stress concentrations which cause creep and slippage in other systems. It is a cost competitive, productive system which should find wide acceptance in coal mines, potash mines, salt mines and essentially any formations in which small boreholes can be drilled.
The DYNA-ROK Anchors
The DYNA-ROK is designed to be used in a nominal one inch diameter hole. Anchorage can be achieved using 5/8", 3/4" or 7/8" diameter roof bolt rods.
ABSTRACT: Finite element method was employed to assess the performance of a two-cavern system over 20-year period of operation. Of primary importance was evaluation of the stability of the pillar separating two caverns. Results of these analyses are discussed along with practical implications relevanto cavern operation.
Utilization of solution-mined underground caverns created within natural salt formations for storage purposes has been increasing steadily in the U.S. over the last several decades. Numerous underground storage sites have been developed and millions of tons of crude oil and other fluid products are stored in solution-mined caverns or abandoned salt mines. After decades of seemingly smooth operation some storage fields exhibit surface subsidence occurring at a rate that concerns the operators with respect to the long-term stability of the cavern systems. It is natural to suspect that high subsidence rates are the primary indicators of potential problems which may arise within an underground structure. Efficient utilization of the salt volume available (i.e. the size of a salt dome) requires that caverns be located at some optimum distance away from each other, such that the operation of one cavern has a negligibleffect on performance of the adjacent cavern. Adjacent caverns are separated by an unexcavated portion of the salt strata, referred to as a pillar. For obvious reasons, this pillar as well as the entire cavern system "should" remain stable over the cavern life. Numerical modeling provides a convenient means by which the performance of underground excavations can be studied. The ever-increasing power of microcomputers permits the numerical modeling of relatively complex systems that in the past were possible only on mainframe computers. The advantage of numerical modeling techniques, such as finite element method, is that the analysis can be performed on the entire system, rather than its components. The purpose of the study discussed in this paper was to evaluate the performance of a simple two-cavern system with particular interest focused on the behavior of the pillar separating two caverns. A finite element method implemented on a microcomputer was used to perform the analysis.
The cavern geometry and location was selected to represent a general case of two caverns of similar shape, located at different depths. A situation often encountered in storage operations in domal salt. The model representing the shape, dimensions, and the relative orientation of the caverns analyzed here is shown in Figure 1. The size of the pillar separating the two caverns was assumed equal to 100 feet. An elastic overburden was assumed to overlay the salt dome to a depth of 1000 feet. A two-dimensional, plane strain finite element model representing a section through the cavern centers was used in the analysis. To limit the time necessary to complete calculations two models were employed: model (a), comprised of 207 4-node isoparametric elements and 167 nodes, and a simplified model (b) comprised of 62 elements and 44 nodes. As indicated earlier, the focus of the analysis was on the stability of the pillar. In the more complex model (a) the pillar is denoted by the following nodes: 54, 55, 56, 57, 58 (top of the pillar), 49, 50, 51, 52, 53 (horizontal mid-section), and 44, 45, 46, 47, 48 (bottom of the pillar).
ABSTRACT: A drill monitoring system has been developed which measures performance characteristics of a rotary drill and relates this to the material strength and rotary specific energy. The drill monitor is computer based and is used to produce a profile of the relative in-situ material strength as a function of depth. This information has been used in determining the placement of explosive in selective loading of blastholes and estimating the amount of lump product in a particular section of an ore body.
In open cut mining, the determination of the location of mineral types and their physical properties generally involves the drilling of boreholes for cores or geophysical logging. However, in many open pit operations there are significant changes in the strata properties over distances of a few metres that cannot be economically characterized through the drilling of exploration holes. The monitoring of the performance of exploration and production blasthole drills has been considered in recent times as a potential alternative to or enhancement of these methods. With this approach detailed in-situ strength information can be obtained without delays to production and the only additional costs involved are the costs of obtaining, installing and maintaining the system. The potential of such a system in tailoring explosive loading to ground conditions in coal overburden has been discussed by Hagan and Reid (1983). The principle behind the drill monitoring technique is that the physical properties of the strata directly affect the behavior of certain aspects of drilling. The relevant drill properties are recorded and the data are used to produce logs of strata strength. The main requirements of such logs is that they reflect changes in strata only and are not influenced by the drill operator. Several researchers have reported studies of drill monitoring. Leighton et al (1982) correlated powder factor requirements with drill performance records in an open pit copper operation. Scheck and Mack (1984) developed a microprocessor based recording system to record drilling parameters and examined correlations between these parameters and cores taken during drilling. Scoble and Peck (1987) have used automated analogue and, more recently, digital recording systems to examine correlations between drill parameters and rock properties. Brown et al (1984) used a prototype drill monitor in exploration drilling. Pfister (1985) and Schneider (1983) have used prototype drill monitors in civil engineering studies.
2 DRILLING PERFORMANCE AND ROCK STRENGTH
Most studies of drilling performance have concentrated on examining factors that affect penetration rate. The following parameters are considered the most significant (Leighton et al 1983, Nelmark 1983):
ABSTRACT: Subsidence may be observed above partial extraction shallow room-and-pillar mines where the coal seam(s) is associated with weak and thick claystone in the floor. This paper presents results of subsidence studies at such a mine. During the 26-month study period, a time dependent subsidence causing a maximum vertical movement of 50 mm was observed. After a subsidence rate of 6.0-6.5 mm/month for the first two months after development of mine workings, the rate decreased to 0.8-0.9 mm/month after 17 months. However, presumably due to settlement of barrier pillars, it increased therafter to 1.7 mm/month for the next 9 months. Angle of draw, based on zero subsidence, was calculated to be about 29 degrees. Maximum tensile strain and compressive strain values observed were 0.040 pct and 0.039 pct respectively. A hypothesis was formulated to correlate surface subsidence movements with observed underground movements. The data collected to date fits reasonably well within the formulated hypothesis.
Trough or sag type surface subsidence may be observed above partial extraction room-and-pillar mines, especially at shallow depths and where the coal seam/s is associated with relatively weak and thick claystone in the floor. An overall extraction ratio of about 50 pct is normally achieved in such a mining system; the remaining coal is left in the form of pillars to support overburden strata. Studies of the characteristics of such subsidence are sparse in the literature. A knowledge of such subsidence is important from a mining as well as an agricultural industry's point of view. It may change land slopes, soil characteristics, agricultural productivity and may cause damage to surface structures. This paper presents results to date of such a study at a mine in the midwestern United States. An attempt has also been made in the study to correlate surface subsidence movements and observed underground pillar settlements as a function of time.
2 MINE DESCRIPTION AND AREA GEOLOGY
The mine extracts a relatively flat 1.5 - 2.0 m thick coal seam at a depth varying from 80 - 100 m. Coal is extracted in panels approximately 1150 m x 250 m (Figure 1) with 5 m wide openings and pillars varying from 17-25 m in different parts of the mine. Barrier pillars of about 54 m width are left between two adjoining panels. The extraction in a panel generally varies 40-45 pct. Roof to floor closure, primarily due to floor heave, is commonly observed in the mine at a long-term rate of approximately 0.5 cm/month. Peak closure rates of 7-10 cm/month are observed immediately after mine openings are developed. The surface topography in the mine area is flat to gently rolling, with a relief of about 6 m. Total thickness of the glacial material over the panel is about 40-45 m while the thickness of the rock overburden over the coal seam is approximately 35 m. The rock overburden primarily consists of shales (65-70 pct), limestones and sandstones. The average overburden thickness over the study panel is 75-80 m. The immediate roof stratum consists of a weak thin band of shale varying in thickness from 0.3 - 2.5 m (average 1.0 m).
ABSTRACT: The portal, which is the near-horizontal, surface point of entry to an underground excavation, can often be an exceedingly difficult area in terms of ground control. Surface and subsurface failures at portals, as discerned during a study involving over 300 case histories, are unfortunately, common. To aid in the engineering evaluation of portals, the Geomechanics Classification System was appended with discontinuity orientation assessment values and support/excavation guidelines. Also, a design model utilizing RMR predicted rock loads in conjunction with half-dome theory is proposed for the most common type of portal failure. Other pertinent comments and recommendations relating to portal stability are included.
The portal zone (Figure 1) is a frequently overlooked and often difficult area in terms of ground control. Failures of rock and support during the 'turning-under process' regularly occur in high angle approach cuts and the initial subsurface portion of the portal. These problems are exacerbated by the weathered, anisotropic, and stress-relieved nature of most near-surface rock masses.
To facilitate portal design for either mining or civil applications, the Geomechanics (RMR) Classification System was appended to aid in the evaluation of the external and internal stability requirements. These modifications are based on a portal database utilizing over 300 case studies (Rogers & Haycocks, 1988), comments from industry designers/contractors, and critical field observations. Furthermore, a design model, based on half-dome theory and the RMR rock load concept, is suggested for the most common type of portal failure. The resulting integration of rock slope and subsurface engineering is necessary for a safe but efficient, long-term portal design.
Since the initial introduction of the Geomechanics (e.g. Rock Mass Rating or RMR) Classification System (Bieniawski, 1973), the concept has been expanded for the engineering evaluation of a variety of specialized areas ranging from dam foundations to rock slopes (Bieniawski, 1984). Current research involving the stability of portals in rock slopes (Rogers & Haycocks; 1988, 1989) acknowledged a need for the rapid evaluation of proposed, existing, and abandoned portals in terms of stability assessment and support requirements.
Figure 1. External and Internal Views of a Portal in a 45°Rock Slope (Rogers & Haycocks , 1989). (available in full paper)
To accomplish this task, information pertinent to portals was gleaned from an extensive literature review, industry contacts, and field investigations. Support system designs, excavation methodologies, and back analyses of portal failures from this database formed the basis for upgrading existing classification system support/excavation guidelines. In particular, it is interesting to note that even though the existing support guidelines from various empirical methods (e.g. RMR, Q, Modified Terzaghi, Corps of Eng., etc.) are considered conservative, data from portal failures indicate that more conservative measures are routinely warranted, especially within approximately I to 5+ diameters inby and outby the portal interface (Barton, et al, 1974; Rose, 1982; U.S. Army Corps of Eng., 1978, 1980).
A shaft is truly the "lifeline" of an underground mine. Damage to the shaft lining and guides as a result of ground movement can result in serious loss of production and extensive and continual repair. In some mining districts, annual shaft repair and maintenance costs resulting from high rock stresses and excessive displacement are in the millions of dollars per year. For many years, the Bureau of Mines has conducted research to characterize the rock mass and to develop structural guidelines to improve the design of accessways in deep mines. Much of this work has centered on determining the in situ conditions that affect the structural stability of shafts in the Coeur d'Alene Mining District of northern Idaho. Recent work by Beus and Board (1984) has concentrated on measuring the rock and support behavior during sinking of the Silver Shaft at the Lucky Friday Mine. These results are currently being used as a comparative data base for further basic research as well as to establish structural design criteria. The factors affecting the design and structural stability of deep mine shafts are numerous. A design approach that uses field data as input and directly compares field measurements with numerical models can validate the procedure and provide realistic design criteria. Naturally occurring "fixed" conditions such as magnitude, direction, and ratio of in situ stresses, geologic environment, and physical properties and constitutive relationships of the rock mass are obviously basic to numerical modeling. Design variables include those factors that have an impact on shaft stability and that may be changed by the designer either before or during construction of the shaft. These are: (1) size, shape, and orientation (if noncircular), (2) type and dimension of support, and (3) excavation and lining sequence. Field data collected from the Silver Shaft verifies the influence of these factors. With present-day technology, it is unrealistic to attempt to incorporate all rock mass behavioral factors and design variables into a single, all-encompassing numerical model. Obviously, some factors have more of an impact on stability than others. The intent of the present effort is to: (1) develop a model that incorporates the effects of opening and support geometries of an actual structure, (2) validate the model by comparing it with actual field measurements, and (3) refine the "displacement reduction factor" in the specific case of the Silver Shaft for use in detailed two-dimensional analyses.
2 FIELD MEASUREMENTS
The Silver Shaft is located about 350 m west of the existing No. 2 Shaft, as shown in Figure 1. Preliminary support of the wall rock was pro- vided by 2-m Split-Set rock bolts and mesh carried to within 1 m of the working face. The liner, nominally at 30 cm thickness, was carried 6 to 10 m behind the face and advanced in 5-m increments. The 6 to 7 m diameter opening was advanced by benching the north and south halves of the shaft at an average rate of 10 ft/d.
ABSTRACT: This paper introduces the concept of design theory and methodology as applied to rock mechanics for more innovative and efficient design. This is a "frontier" research area in rock mechanics because although design is a fundamental foundation to all engineering branches, very little attention has been paid to this aspect in mining. In discussing the theoretical premise of design, this paper identifies the principles which form the basis of design activity. Also, from a practical viewpoint, the findings from a series of interviews with mine design engineers are presented. These findings lead to the identification of needs arising from overlooked design methodology in practice.
In recent years, design theory and methodology have begun to emerge into a new discipline which supports all fields of engineering. However, concepts of design theory and methodology have neither been applied nor systematically studied in the field of mining engineering. This paper introduces these concepts as applied to rock mechanics for more innovative and efficient design, thus establishing a "frontier" research area. The importance of design research in the field of rock engineering is becoming increasingly evident. While it is true that there have been many impressive achievements in mining over the last several years, it is also true that innovations in mining have occurred at a much slower rate compared to other engineering disciplines. For example, it has been observed that the introduction of rock bolts in the 1940's represented the last major innovation in the area of strata control. Still today the layout of rock bolts is based primarily on empirical procedures. In summary, research into rock mechanics design is justified as follows:
ABSTRACT: Coal strength based on scaled uniaxial compressive strength from the laboratory and back-calculations from in-mine observations of pillar stability indicate that material strength scaling rules are open to debate. Analyses presented indicate that the application of scaling factors lower than 0.5 provide a better correlation with field observations of pillar stability.
Pillar sizing based on empirically derived formulae has been performed since the early 1900's. The current art involves the scaling of laboratory-derived coal strength to values representative of the in-situ coal mass. The various pillar-strength formulae, incorporating conventionally accepted scaling concepts, were applied to size pillars at an underground coal mine located in Pennsylvania. Even for presently stable pillars at the mine, the calculated safety factors were below those deemed adequate. Back-calculations based on observed pillar instabilities yielded an in-situ coal strength ranging from 730 psi (5.0 MPa) to 1073 psi (7.4 MPa) depending on the particular pillar-strength formula utilized. These values are approximately of the same magnitude as the coal strength determined in the laboratory [700 psi (4.8 MPa) to 1500 psi (10.3 MPa)]. As a consequence, the application of published scaling relationships to these laboratory data produced predicted in-situ coal strengths much lower than those that apparently exist based on the observed in-mine conditions. In this paper, a series of typical steps leading to determination of a "safe/stable" pillar size are applied. Discussion includes details relevant to the laboratory as well as field investigations, that generated concerns related to the use of common pillar formulae.
SPECIMEN PREPARATION AND TESTING
Although sample material for three coal sems (referred to as S-1, S-2 and S-3) was prepared and tested, it should be noted that the pillar sizing discussion will ultimately refer to only one of these seams, namely seam S-3. The samples for all three seams, as received, averaged one foot wide, two feet long, and one foot thick (across bedding). The sample material was obtained from recently mined area and delivered to the Penn State Rock Mechanics Laboratory wrapped in heavy plastic bags and burlap.
Table I, Results of Uniaxial Compressive Strength Tests on Coal(available in full paper)
Preparation of Specimens
The lump samples were kept moist and wrapped until specimen preparation, since it was expected that oxidation and progressive cracking from the drying of the coal material would deteriorate the structural integrity of the lump severely. Initially, specimen preparation efforts concentrated on obtaining cylindrical specimens using a standard laboratory coring techniques. It became apparent, however, that the specimen preparation techniques had to be altered for coal material from each seam. The standard coring technique provided useable test specimens for seam S-1 only. The material from seams S-2 and S-3 was very friable and fractured readily as coring was performed. Consequently, cubical test specimens were prepared from seam S-2 and S-3 material, exclusively. A horizontal band saw equipped with a tungsten-carbide blade was used to successfully prepare these cubical specimens. The ends of all specimens (cylindrical and cubes) were finished by hand with 200-grit sandpaper. Following preparation, specimens were kept moist in an environmental chamber (100% humidity).