RESUME: La rupture des roches autour des ouvrages souterrains est un phenomène courant à grande profondeur. Dans le domaine minier, ces phenomènes se traduisent, dans la majorite des cas, par des deformations et des deplacements d'ensemble des ouvrages qui, pour etre parfois très irnportants, se manifestent de manière relativement continue, tout au moins à l'echelle macroscopique où on les observe. Toutefois, dans certains secteurs, ces ruptures prennent des formes singulières par leur brutalite et l'ampleur de la zone affectee: ce sont les coups de terrains ou "rockbursts" de la terminologie anglo-saxonne. A partir de l'examen des differents cas observes dans les mines de charbon françaises, l'article precise les facteurs qui apparaissent jouer un rôle dans le declenchement des coups de terrain. Cette approche propose des mecanismes de rupture possibles à l'origine de ces phenomènes, ce qui permet de rendre compte et d'expliquer, dans une certaine mesure, leur caractère discontinu. Les coups de terrains se manifestent par des ruptures brutales du massif rochcux auteur des cavitues minières. Par rapport aux phenomènes de ruptures plus classiques et frequemment observes aux profondeurs d'exploitation actuelles, ces manifestations s'individualisent par leur soudainete, leur violence et, d'une manière generale, par l'etendue de la zone affectee. Dans les mines françaiscs, mis à part les phenomènes d'effondrements spontanes observes dans certaines exploitations conduites par chambres et piliers, c'est surtout dans les mines de charbon que ces phenomènes se sont manifestes avec le plus de diversite au cours des dernières annees. Actuellement, on peut avancer que 25 à 30 % de la production nationale de charbon sont issus de secteurs susceptibles de presenter des coups de terrains au cours de leur exploitation. 1 LES COUPS DE TERRAINS - LEUR ENVIRONNEMENT 1.1 Les phenomènes observes Dans les mines de charbon françaises, les coups de terrains observes depuis une dizaine d'annees concernent les exploitations par longues tailles. Les effets notes au niveau des chantiers (fig. la) consistent en une rupture et une expulsion de la couchc de charbon (coup de couche), accompagnee d'un ebranlement violent qui peut etre ressenti jusqu'en surface (JOSIEN, 1981). La magnitude de cette secousse a pu être estimee à partir des reseaux sismiques installes: elle peut atteindre 3 pour les plus importantes. Dans ce dernier cas, de grandes longueurs peuvent être affectees (superieures à 100 m), ct le soutènement porteur de la galerie est chasse ou renverse, et même ecrase si le deplacement impose n'est pas possible. En ce qui concerne lc type d'exploitation, differents cas ont ete rencontres: - veines de charbon en "plateures" (pendange inferieur à 20°), exploitôes par foudroyage (fig. 1a) - veines de "semi-dressants" (pendage compris entre 20 et 50°), foudroyees ou remblayees (fig. 1b) - veines en "dressants" (pendage superieur à 50°), exploitees par tranches montantes remblayèes (fig. 1c). 1.2 Caracteristiques des charmers affectes Les exploitations par tailles sont affectees par ces phenomènes depuis une dizaine d'annee et, au cours de cette periode, une soixantaine de coups de terrains d'importance variable a ete notee.

ABSTRACT In present work the systematic investigations on the mechanical characteristics of rocks under tree-dimensionally stressed state conditions, close to those, acting within the massif, are described. The program on mechanical tests of rocks is carried out, depending upon the calculated qualitative changes of principal stresses and three-dimensionally stressed states near the mine workings. The developed energy criterion of rock strength includes the deformation forces, acting under the deformation conditions, the shape and volume changing, their limiting values and the Nadai-Laudet parameter, taking into account three-dimensionally stressed state types. 1 INTRODUCTION The rock and rock massives loading conditions have a pronounced dimensional character. The relations between three principal stresses, characterizing mechanical state of rock massif near the mine workings are changed naturally. It defines the corresponding massif deformations and failure. The investigating the complex rock massif deformation and failure processes would be more successful under the natural conditions. However, the single tests only are possible in this case because of substantial production problems. The thorough preparation work for mine tests includes the necessary laboratory and analytical study. In present work the systematic investigations on the mechanical characteristics of rocks under the three-dimensionally stressed state conditions, close to that, acting within the massif, are described. this study allows to reveal the most characteristic features of geo-mechanical processes, proceeding Within the massif and to determine the main process parameters changing trends. 2 VOLUME LOADING OF THE MASSIF As for the gentle sloping rock bed in front of the face, all three principal stresses, acting in geo-mechanical processes, which proceed in the above mentioned zone within mine workings, are compressive, and three-dimensionally stressed state in different massif areas varies from summarized compression to summarized tension (Norel 1983). Three principal stresses, defining the three-dimensionally stressed state within the massif, are varied qualitatively as follows. When approaching the face, the maximum compressive stress, directed vertically relative to the working plane and defined as б3, increases from the vertical compressive stress value within unmined massif zone up to the peak value, which equals in the maximum compressive stress zone (where Kk is the concentration index) and then falls down to the strength index value at the face edge. The intermediate compressive stress, directed over the length of the face and defined as б1, varies from λp3 within unmined massif zone to the 1/2 (б1+ б3) value in maximum compressive stresses. 3 TESTS PROGRAM The program on mechanical tests of rocks is carried out depending upon the calculated qualitative changes of principal stresses and three-dimensionally stressed states near the mine workings. Before analyzing the test results, some statements on deformed media mechanics. being used in this work (Nodai 1969) should be considered because there is always the algebraic inquality б1≥б2≥б3 for three triaxial principal stresses. In this case the compressive stresses and deformations are negative; tensile stresses. 4 TENSOR DECOMPOSITION Tensor decomposition for three main types of three-dimensionally stressed states has allowed the more clear understanding of physical foundations of the summarized compression.

ABSTRACT: The effect of injection-fluid flow through the borehole wall on hydraulic fracture breakdown pressure was studied theoretically and in the laboratory, for both low and high permeability rocks, under initially dry or saturated (under pore pressure) conditions. The results indicate that the conventional Hubbert and Willis equation developed for fluids that do not infiltrate rock does not apply, and suggest that the poroelastic model with appropriately derived Biot parameter and Poisson's ratio is more realistic. Fortunately, the difference between the two approaches is not significant under typical in situ stresses at shallow depths. However, at great depths (say over 1–2 km) use of the proper stress-pressure relation is critical if a more realistic maximum horizontal stress magnitude is to be determined. 1 INTRODUCTION Wells deeper than ever before are being drilled for oil production, for hot dry rock geothermal energy, for waste disposal, and for scientific purposes. The Kola Peninsula scientific well in the Soviet Union has reached the unprecedented depth of over 12 km, and ultra deep scientific borings are underway in Germany, Sweden, and elsewhere. However, most of our hydraulic fracturing (HF) stress measurements have been limited to shallower depths. Outside of the Michigan Deep-Well tests (Haimson, 1978) in which measurements were conducted to a depth of 5 km, hydraulic fracturing has so far been confined to the top 2 km below the surface. A major problem with deeper wells is that rock stresses tend to become larger, affecting the stability of the borehole with respect to both compressive failure as expressed in the form of breakouts, and tensile failure when subjected to internal pressurization as in hydraulic fracturing. At the University of Wisconsin, we have been investigating in the laboratory the mechanical conditions under which wellbore breakouts occur (Haimson and Herrick, 1986, 1989), and hydraulic fractures are induced (Haimson and Edl, 1972; Edl, 1973; Haimson and Huang, 1988; Huang, 1989), for in situ stresses ranging to levels encountered in ultra deep holes. This contribution summarizes our findings related to the stress-pressure relation in deep-well HF stress measurements. 2 THEORETICAL CONSIDERATIONS The conventional elastic (E) model used to calculate in situ stresses from HF pressures assumes that the rock surrounding the wellbore is linearly elastic, isotropic, and impermeable to the fluid injected into the borehole (Hubbert and Willis, 1957). They used the known solutions of the "hole in a plate" and the "hollow thick cylinder" problems in linear elasticity to establish the relation between the tangential stress at the wellbore wall St, the injection-fluid pressure and the far-field minimum and maximum horizontal stresses Sh and Sh. They asserted that hydraulic fracturing is initiated when St becomes so small, as a result of increasing injection-fluid pressure, that tensile failure occurs. The criterion used was Steff- -T (where T is the tensile strength of the rock, and Steff- St - Po; Po is the pore pressure, and Steff is the effective minimum tangential stress at the wellbore wall, related to St by Terzaghi's effective stress law).

Abstract: Rock masses, which commonly contain a large number of discontinuities like joints, are treated as homogeneous, anisotropic porous media. The corresponding permeability tensor k, is formulated in terms of a symmetric, second-ran k tensor pij, which is dependent not only on the geometry of related joints (density, size, orientation and aperture of joints) but also on the earth pressure at any depth. Two case studies are carried out; that is, the ventilation drift at Stripa Mine, Sweden, and the underground cavern for an oil storage at Kikuma Test Plant, Japan, where the large scale hydraulic conductivity tests have been done together with the extensive survey of joints. It becomes clear that a statistical interpretation on the joint data collected from an in-situ rock mass provides a powerful tool to determine the permeability tensor. The mean permeabilities of the both sites accord well with the theoretically predicted ones. On the basis of this study, the permeability at great depth is considered. 1 INTRODUCTION In the recent topics of rock hydraulics, much attention is focused on ground water flow through various geological discontinuities to solve some problems concerning the geothermal energy, earthquake and deep underground burial of high level nuclear waste. For example, the repository of high level nuclear waste will be buried about 300 m to 1000 m below the surface in rock masses (e.g. Runchal and Maini, 1980). One of the most serious problems to be solved is the isolation of the high level nuclear waste form the biosphere. Ground water flow through geological discontinuities (called cracks) is believed to be the most significant way of radionuclide migration. Field tests can be a practical solution to overcome the present difficulty. Large scale hydraulic conductivity tests were carried out to investigate the hydraulic properties of a low-permeable, jointed granue, for example, at Stripa mine in Sweden (Wilson, Witherspoon, Long, Galbraith, Dubois and Mcpherson,]983) and at Kikuma in Japan (Hoshino, 1983 and 1984). In addition to such field tests, many theoretical (or numerical) studies were also done to give a sound basis for predicting the overall permeability of rock masses; e.g., Parsons (1966), Snow (1969), Wilson and Witherspoon (1974), Long, Remer, Wilson and Witherspoon (1982), and Robinson (]984). Among others, Oda, Hatsuyama and Ohnishi (1987) have also proposed a theory in which discontinuous rock masses are treated as homogeneous, anisotropic porous media. By extending the previous study, we will propose a set of equations which make it possible to predict the permeability of jointed rock masses at great depth. 2 HYDRAULIC PROPERTIES OF DISCONTINUOUS ROCK MASSES 2.1 Permeability tensor In order to formulate a permeability tensor for discontinuous rock masses, the following idealization was accepted (Oda, et al., 1987): Any crack has a shape similar to a penny, with a diameter r and a uniform aperture t, whose orientation is identified by a unit vector n normal to the principal plane. Her, r, t and n are random variables.

ABSTRACT: Fluid-rock interactions control the evolution of pores and cracks. At geological time scales, two important physico-chemical processes of deformation are subcritical crack growth and pressure solution. The first of these can explain various observations such as time dependence of strength, anelastic strain relaxation of cores, etc. The second process is a key one as far as the porosity structure is concerned through the crust. 1 INTRODUCTION Crustal rocks are made of a matrix and voids. These voids are either pores or cracks. Fluids are circulating through the connected porous network. During this circulation, they interact with the rock matrix. Thus two questions arise, which are: what are these interactions ? and what role do they play in the evolution of the void structure ? In fact, we can define two types of interactions. First, there is a purely mechanical interaction hydraulic fracturing develops when the increasing fluid pressure overcomes a critical value (Charlez et al., 1989, Cleary, 1989). On the other hand, fluidrock interactions can be of a physicochemical nature and lead to two long-term evolution processes: slow crack growth and pressure solution. Both of these phenomena are active on a time scale which is larger than the laboratory scale. In this paper, we focus our attention on slow crack growth and pressure solution. We present both of these processes and we give for each of them consequences and applications which are of great potential interest for those who are dealing with the long term stability of reservoirs, or man-made underground constructions. 2 SLOW CRACK GROWTH 2.1 What is slow crack growth? Let us consider an elastic solid containing a crack and submitted to tensile stresses. Following Griffith's (1920) energy criterion, the system (solid + crack + loading device) is at equilibrium when its energy is at a minimum. When G > Gc, the crack propagates dynamically (velocity greater than 1 m/s). On the opposite, when G < Gc, the crack is expected to remain stable, or even close if only reversible phenomena are taking place at the crack tip. However, numerous studies on different materials (ceramics, glass, single crystals, rocks) have shown the existence of slow crack growth or subcritical rack growth (velocities lower than 10- 3 m/s) for G < Gc (Wiederhorn, 1968; Atkinson and Meredith, 1981; Swanson, 1984; Reuschle et al., 1989a). These observations can be explained by the existence, in real materials, of irreversible phenomena (plasticity, viscoelasticity, frictions,…) at the crack tip (Maugis, 1985). Because of these internal frictions, the critical crack extension force for dynamic crack propagation (Gc = 2 Yc) is greater than the crack extension force necessary to maintain the crack at equilibrium. This last parameter is G0 = 2 γ0' where γ0 is the thermodynamic surface energy (in vacuum).. A clear distinction has to be made between γ0 and γc: γ0 is a thermodynamic surface energy which takes into account reversible phenomena; γc is the surface fracture energy which takes into account irreversible phenomena.

Introduction This Symposium General Report attempts to highlight important developments in Rock Mechanics for applications to engineering problems at great depth. Much of the report was presented orally in session # 9, but in an abbreviated form. The many excellent written and oral contributions to the Symposium represent a technical richness that will take time to assimilate. We cannot mention all the technical content nor do justice to the effort that the Proceedings represent. We recommend that the reader scan the various General Reports and Invited papers to appreciate the breadth of technical content and the scope of the topic "Rock at Great Depth". Then, with this as background, the detailed papers will provide insight into a number of relevant Rock Mechanics problems in the areas of civil, mining and petroleum engineering, with links to geochemistry, solid body mechanics, geo-hydrology, applied mathematics, tectonics and geophysics. The interdisciplinary scope of the papers is astounding; when a mine designer and a geochemist sit down and discover common interests, when a drilling engineer and an applied mathematician outline similar views of a phenomenon, it speaks for the excitement of Rock Mechanics applications at great depth. The Session General Reporters and Workshop Presidents provided us with useful oral and handwritten comments. In particular, we thank A. Nur, D. Ortlepp, J. Geertsma, M. Chenevert, M. King, R. Widman, W. Bradley, C. Spiers, and O. Coussy for their opinions and comments. Conclusions from the Workshops are summarised, based on oral summaries provided by A. Guenot, W. Kessels, F. Cornet, and written comments by J. Sulem, I. Vardoulakis, P. Berest and Ph. Weber. The last-minute communication by B. Come has also been used. Before a general Symposium conclusion is attempted, we will address a number of current and central issues in Rock Mechanics – Unique aspects of the mechanics of geo-materials at great depth – Rapidly developing areas related to Rock Mechanics and – Specific issues that have arisen during workshops. 1.0 - UNIQUE ASPECTS OF ROCK BEHAVIOR AT GREAT DEPTH "Deep" is defined as 3 000 to 7 000 m, occasionally somewhat deeper, but still involving situations where the rock is physically encountered, as opposed to remotely probed. Below 10,000 m, one must speak of "geological" or "geophysical" depths. This may be the domain of the next generation of petroleum and rock engineers and applied earth scientists; currently, only the deep Kola Peninsula scientific borehole exceeds this limit. A number of special and important cases were addressed during the technical presentations. 1.1 - Rocks with Abnormal High Porosity with respect to Depth Exceptionally high porosities at depth are found in chalk, silt, and sandstone reservoirs as well as undercompacted shales. Examples include the Chalk reservoir of the Valhall Field in the North Sea, with beds of 50% porosity at 3 000 m depth; the "geo-pressured" reservoirs of the Gulf Coast, California, and many other basins; and high porosity shale cap rocks in many areas, usually with pore pressures well above hydrostatic.

ABSTRACT: The elastoviscoplastic behavior of saltrocks may be interpreted in terms of a number of different fundamental deformation mechanisms that act at different rates at different shear and normal stress levels. The paper presents a new approach to a constitutive law for these materials, recognising the complex behavior, and making distinctions between different types of yield criteria and viscous flow laws. An attempt is made to develop this behavior as an analogue to the behavior of rock at high temperatures and stresses, where viscous mechanisms became possible. It is empasised that an appreciation of the dominant deformation mechanisms acting at particular conditions will guide the form of macroscopic constitutive laws for non-saltrocks. INTRODUCTION: Understanding rock behavior at great depth and elevated temperature requires quantification of deformation processes which do not occur in ordinary conditions. Study of these time-dependent (creep) processes requires high temperature (T) and stress (σm¹) tests at slow strain rates (dE/dt). Shear stresses (r=(σ¹-σ 3 )/2) and pore pressures must also be controlled. It may be of practical interest to explore analogues which display similar mechanisms at lower T, σm¹, and r conditions. Salt and potash may be useful analogues in some conditions. Halite (NaCl) and sylvite (KC1) display different deformation processes depending on σm¹, T, and r, but in conditions accessible in the laboratory and the field (σm¹ <70MPa; r<40MPa; T<150°C). Provided that it can be shown that mechanisms are similar, analogies may be developed to other rock types. The critical issues are that behavioral fields, activation energies, and mechanism transition points be quantified to permit comparisons and extrapolations; deformation maps (Frost and Ashby, 1982) are essential for this purpose. A behavioral model for saltrocks will be described and treated as an analogue to the behavior of rocks at great depth. Emphasis is on the effects of stress and strain rate on mechanisms in saltrocks. Saltrocks display most aspects of brittle viscous- ductile behavior at strain rates of d /dt=10- 2 -10-¹ 2 S-¹: brittle yield and dilation at low σm, ideally elastoplastic and non-dilative behavior at high σm' transition from viscoelastic to viscoplastic creep as r rises, of steady-state cracking (with an-nealing), rate dependence of the yield criterion, and primary creep with recovery. YIELD BEHAVIOR: Elastoplastic (EP) yield is the conventional yield criterion (Y) of rock mechanics. Fig. 1 shows Brazilian and triaxial test data giving a curved criterion with these features: at low σ 3 , brittle behavior with dilatancy and debonding, small to failure, dependency of Y on σ3, UCS about 12 times To; at high σ3, cataclastic flow without σ3 dependence, dilatancy supression, crystal rupture; and, at intermediate confining stresses, transitional behavior. Typically, for halite, the transition to fully cataclastic plastic yield occurs at σ3= 10–20 MPa, σ1-σ3]EP≈ 35- 40 MPa, with rate effects. Consider a test at low σ3: if r<Y, steady state creep will occur, but r>Y can be applied for short times before strain-weakening. Thus, Y is a lower bound, defined as the lowest τ for strain-weakening yield in any time.

ABSTRACT: This General Report requested by the Organisers attempts to describe the state-of-the-art as perceived by the Reporter in the area of the determination of rock properties at great depth by physical methods. The paper focuses on laboratory studies of rocks under physical conditions simulating those crustal conditions in which a rock may be modelled as a cracked solid through which fluids may percolate. Current and upcoming developments in the study of compaction and dilatancy in such rocks under triaxial loading are summarised, and the intimate relation between rock mechanics and rock physics is brought out. An important special aspect of this study is the transition from cataclastic to plastic behaviour, which coincides with the elimination of dilatancy and probably of seismic activity. The paper also discusses the need for a strong parallel focus on crustal stress measurement, which is the missing link in the application of rock mechanics and rock physics to many engineering and Earth science problems, particularly in the context of plate tectonics theory, which has facilitated the systematic ordering of knowledge about the Earth's crust and its properties, processes, and evolution. 1. INTRODUCTION With increasing knowledge of the origin, processes, and properties of the Earth's crust, and increasing engineering and technological capacity to extract minerals and heat energy from great depths, there is naturally a growing desire to measure and understand the properties of rocks and rock masses at these depths. The concepts and theories of plate tectonics have in recent years enabled our knowledge of the crust to be ordered in a fundamental and systematic way more suited to useful physical and chemical modelling (e.g. see Bott 1982, Turcotte & Schubert 1982, Meissner 1986). This has stimulated a new interest in the nature, processes, origin and properties of the lower parts of the continental crust (e.g. see Dawson et al 1986), which are being increasingly probed, by electrical deep sounding methods (Haak & Hutton 1986), deep seismic reflection profiling (Meissner 1986, Matthews 1986, Smithson 1986), (which provide both rock samples and opportunities for close range measurements at great depths, see Raleigh 1985 and Behr et al 1987). Field studies by geophysical and geological methods enable the properties of the crust and of rock masses within the crust to be determined and measured, but this is dependent on a good understanding of rock properties at the crystal, grain, and atomic scales. The latter properties have to be determined in the laboratory, based on the laws of physics and chemistry. The subject of the present paper is the determination of rock physical properties in the laboratory within the perspectives of this Symposium. 2. LABORATORY MEASUREMENTS OF ROCK PROPERTIES Advances in laboratory techniques are enabling measurements to be made of rock physical and chemical properties under simulated crustal conditions corresponding to increasingly greater depths. Indeed it is now possible to simulate most crustal conditions, though there are very few laboratories at present that are fully equipped for measurements on rock samples under simulated conditions of great depth.

ABSTRACT: At great depth stresses are approximately equal to the strength of rock. That is why failure develops near such underground excavations. But this failure may not be entire because its value depends not only from stresses but also depends from strains. On the basis of experiments on compression specimens of rocks in rigid presses, particular on decreasing the cohesion with increasing the strains it was formulated the model of gradual shear failure of rocks. Models with constant parameters in the Coulomb formule for the cohesion do not reflect the real decreasing of cohesion and lead to diminish displacements of rock near underground openings. On other case models with instant decreasing the cohesion leads to essential greater displacements because unjustified diminish the cohesion in calculations. It was revealed when the model has been used to one-dimensional static problem of compression rock with circular hole. Also it has been calculated the dynamic problem of burst in rock. Another application of the model has been made to two-dimensional problem about influence the gradual failure of rock into geological break or dyke, which appear near underground opening on formation dangerous vertical tensile stresses. THE MODEL OF GRADUAL SHEAR FAILURE OF ROCKS Rocks contains many strucktural defects, holes and cracks. If stresses change through the distance much larger than the interval between structural defects, then such rocks may be regarded as a continuous medium. The main effect of structural defects and cracks consists in two main phenomena: in decreasing of a cohesion and in increasing of a crack porosity. These two phenomenon a develop simultaneously but gradually while the nonreversible shear strains increases. It follows from numerous experiments of Karman (1911),Byerlee (1968),Stavrogin and Protosenya (1985) and others, and have been used in developed by author the model of gradual shear failure of rocks (Korotkov (1980,1987,1989)).. The model consists of several main equations. In the beginning of failure of rocks there are some strengthening during small increasing of nonrevercible shear strains; it defines as prefailure. Further increasing of nonlinear shear strains causes the intensive decreasing of shear strength. This region defines as post failure. These effects may he described by the loading surface depending from nonreversible shear strains also pressure and temperature. The value of YP calculated only in the region of failure where pressure is not too high, usually less then 10 kbar, because at more high pressure the rock deforms without loss of compressive strength. The full volume is divided in two parts. It consists reversible part of solid framework and nonreversible volume of crack porosity. For the framework the equation of state was accepted with the consideration of effective value of pressure. Earlier theories used only one-term for description of nonreversible porosity, such as Rudnicki and Rise (1975) included only one term for description of dilatancy in rock. Other theories for granular material also used one term for description the dependence porosity from pressure. The model of gradual shear failure used more general description with including both mentioned terms.

RESUME: L'instabilite à grande profondeur des massifs rocheux au voisinage des vides sou terrains est liee au developpernent des zones fracturees par suite d'un exces de contraintes. La modelisation du comportement post-rupture en termes de plasticite avec ecrouissage radoucissant permet de simuler, soit analytiquement, soit numeriquement, le developpement de cette fracturation mais cette modèlisation doit être amelioree pour rendre compte du caractère explosif de certaines instabilites de la localisation de certaines ruptures et de la cinetique de la propagation des fractures. 1. INTRODUCTION Il convient, en preambule, de souligner toute la relativite de la notion de "grande" profondeur dans le contexte qui nous occupe; en effet, pour l'ingenieur mecanicien des roches, la profondeur est grande si les contraintes qui prennent naissance au voisin age des vides qu'il cree excèdent la resistance de la roche dans laquelle ces vides ont ete realises. Ainsi, dans les communications etudiees par le rapporteur, la "grande" profondeur se situe-t-elle à 230 m, 800 m ou 3500 m selon que les auteurs ont affaire à l'argile de Boom, au charbon de Coventry ou aux quartzites aurifères d' Afrique du Sud. II. FORMULATION ANALYTIQUE DU CONCEPT DE RUPTURE PROGRESSIVE. La caracteristique des excavations souterraines à grande profondeur etant par definition l'existence de ruptures à leur voisinage, l'un des principaux progrès realise par la mecanique des roches pendant ces vingt dernières annees a ete de comprendre la necessite d'une modèlisation du comportement mecanique des roches fracturees. Dans les annees 1970 les premiers resultats experimentaux ont ete publies (WAVERSIK (1968), WAVERSIK et FAIRHURST (1970), RUMMEL et FAIRHURST (197)) qui montraient que la rupture des roches pouvait ne pas être explosive et que la perte de resistance pouvait être progressive à condition de disposer d'une presse asservie suffisamment rigide. On a ainsi vu apparaître la notion de comportement post-rupture qui allait permettre de mieux comprendre les problèmes de stabilite et de soutènement. P.EGGER en 1973 a ete l'auteur de la première formulation theorique du comportement post rupture et de la perte progressive de resistance en termes de plasticite à ecrouissage negatif ou radoucissant. Ce modèle lui a permis d'etudier l'interaction, entre massif rocheux et soutènement et de jeter les bases d'une methode de dimensionnement des revêternents de tunnel qui allait acquerir une grande notoriete sous le nom de "convergence confinement" Ulterieurement PANET (1976), BEREST, NGUYEN MINH DUC et PANET (1978) ont precise le concept d'ecrouissage radoucissant en formulant de manière rigoureuse la loi de comportement du milieu continu elastoplastique equivalent au massif rocheux fracture. Les calculs analytiques qu'ils ont developpes s'appliquent au cas de tunnels circulaires et de cavites spheriques et supposent un champ de contraintes initiales isotrope. On retrouve dans le present symposium ce type d'analyse dans les communications à caractère theorique de P.F. KOROTKOY d'une part, CHEN JIN, YUAN WENBO et CHEN XIANGFU d'autre part. Un modèle de comportement legèrement different est propose par HISATAKE CORDING, ITO, SAKURAI et PHIEN-WEJA avec le souci de representer de façon plus realiste le comportement du massif rocheux.

ABSTRACT: A borehole radar system with a range exceeding 100 m in granitic rock has been developed and tested within the Stripa nuclear waste project. This system has been applied in measuring sites all over the world and can presently be used down to 1 km. Different measurement techniques are described: in particular the recent development of a directional antenna makes the system useful in deep boreholes. 1. INTRODUCTION Radar methods have been applied in geological investigations for about twenty years. The pioneering work was performed by Cook, who studied the electrical properties of rocks, soils and clays at different frequencies (Cook 1975). He demonstrated that electromagnetic waves can produce radar pictures in a limited frequency window where the range will be some tenth of meters into the ground. Several commercial system have been constructed in the frequency range 100–1000 MHz. These instruments have been successful in locating pipes, measuring the thickness of an overburden, investigating roads and building sites etc. Ground radar has sometimes suffered from a poor reputation caused by over optimistic practitioners who have neglected the difficulties of interpreting the radar data: geological structures are in fact often too complex to admit a unique interpretation, but if these problems are properly taken into account the ground radar can be an efficient instrument at small depths. 2. BOREHOLE RADAR The limited range of a ground radar makes it highly desirable to measure with radar from boreholes in order to cover a larger rock volume. Core mapping and borehole geophysical logging can provide information about the quality of rock but are sensitive to the very limited sampling provided by the borehole. Most logging instruments are sensitive to properties only a few cm from the borehole walls and it is not uncommon that important structures are missed because the borehole was misplaced. The risk can be reduced by drilling many boreholes in the same area, but if the borehole is very deep this method is excluded for economical reasons. A borehole radar can alter the situation radically, since the range of this instrument can be more than 100 m in crystalline rock. The range is mainly determined by the water in the rock. Dense crystalline rock will thus be more penetrable the porous rocks though radar measurements are possible even in sedimentary rock. Apart from range a logging instrument must also provide good resolution. The total amount of information collected with a borehole radar surpasses most other logging instruments, since the radar can produce hundreds of values at each measurement point in the borehole. High resolution requires a short pulse. In practice the pulse length is nearly equal to the wave length, which means that the pulse is very broadband in frequency. The resolution is, however, even better than a wavelength, since the system has a time resolution which can resolve details inside the pulse. The wavelengths in rock are 2–5 m with the RAMAC system, but reflectors can be located with a precision better than 0.5 m.

ABSTRACT: It is necessary to grasp joint distribution in rock masses and modelling it for reliable and effective analyses of groundwater flows. Joint information which could be obtained from the site is limited because of different conditions of site investigations. Therefore, statistical plans to estimate the joint distribution parameter must be discussed for a parametric estimation. We suggest serial job flows for joint sampling, estimation, and modelling methods (2-D, 3-D) of joint distribution through limited information to be obtained at sampling points. 1. FOREWORD Recently, with the increase of constructions of large underground structures, studies concerned with rock joint distribution, which will affect foundations at great depth, have been made actively, though, there are only few studies on development of modelling systems of jointed rock masses as a serial system to be integrated whole procedures of surveys, statistics, and modelling as a stage now. Therefore, the purpose of this study is to make a model of jointed rock masses in order to integrate the results of studies1)2)3) since before on characteristic elements of joint distribution. Further, we tried to formulate to systematize the results by introducing new methods, and developed a calculation system for modelling. 2. OUTLINE OF THE SYSTEM In this system, in order to consider the local deviations of regularity of joints, we classified the objected area into regions according to the distributional characteristics being statistically judged as a similar one, and we clarified the joint population in the objected area into several partial populations (joint sets) as clusters through judges of special similarity of joint distribution according to the orientation of joints. Then, based on a probabilistic space model defining each characteristic element such as 1) orientation 2) persistence 3) density 4) aperture 5) filling 6) roughness 7) conductivity by each joint set, we recognized probabilistic structure through stochastical point estimation or interval estimation of parameter to be necessary for expressions which using functions. Moreover, we made a model by means of simulating the joint system based on the probabilistic space. (Fig. 1) 3. PROBABILISTIC SPACE DEFINITION OF JOINT DISTRIBUTION1) For the purpose of making a stochastical simulation model of joint distribution, probablistic structure of each characteristic elements of joint distribution shall be defined. 4. SAMPLING SURVEY OF JOINT DISTRIBUTION Generally, when considering that the number of distributed joints is large and survey able area is limited, it is natural that joint distribution survey has been carried out by sampling surveys. However, the results of sampling surveys shall be apply able to estimate the probabilistic structure which shall be defined as abovementioned clause 2. As for this kind of sampling survey method, it is a general method since before to set scan lines or scan windows on the surface of rock masses and to describe about the crossed joints which is longer than the certain defined length (Q.,). Further, in order to set sampling points, it will be preferable to select the points according to the homogeneous sampling method of which work efficiency is high.

ABSTRACT: Geomechanical parameters of in situ rock masses, including rock stress, deformation modulus, Poisson's ratio, and strength have been measured or determined simultaneously utilizing a combination of three U.S. Bureau of Mines hydraulic borehole pressure cells installed in one hole. Two case studies were conducted at a western U.S. coal mine in two deep-seated and adjacent seams under high horizontal stresses. Measured vertical rock stresses generally agree with the theoretical values derived from the respective overburden loads. Deformation modulus is highly stress dependent, and the plot of transient deformation modulus versus normalized axial stress increment is a parabolic function of the form y=ax 2 , up to the yield point. Poisson's ratio decreases during the initial loading period, then stabilizes with small fluctuations until failure. Coal strength increases in accordance with increasing load and confinement even beyond the yield point and up to ultimate failure. Failure strength under highly confined conditions is substantially higher than the unconfined strength. The failure strength is a function of the intermediate confining stress, as well as of the minimum confining stress. 1 INTRODUCTION Knowledge of the geomechanical parameters of rock masses is needed for both design and stability assessment of the underground excavation. These parameters include, among others, the state of stress and the strength and deformability of the rock mass. Strength is of particular importance if underground excavations are located at great depth or exposed to high confining stress fields for which the failure mechanisms are stress-induced. Except in a few instances, (Mogi, 1971, 1972; Amadei and Robison, 1986; Lu, 1988) the effects of general triaxial stress states on the in situ strength of rock masses have not been extensively studied because of the complexities and difficulties in determining general or true triaxial strength. In order to assess the stress-induced instability of an underground structure, the rock mass strength must be determined under in situ loading conditions or, if determined in the laboratory, under simulated in situ conditions. This procedure was suggested by Lu (1986a) to assess the three-dimensional stability of mine pillars with the "integrity factor" model. For rock stress determination, four methods, "Flat Jack," "Hydraulic Fracturing," "USBM Deformation Gage," and "CSIR or CSIRO Cell", have been recently suggested by the ISRM Commission of Testing Methods (Kim and Franklin, Joint Coordinators, 1987). However, besides the fact that they are indirect stress measuring methods, their field operations are tedious and expensive. For rock deformability determination alone, the ISRM Commission on Testing Methods have also suggested two methods: "Large Flat Jack" technique (Loureiro-Pinto, J., Coordinator, 1986) and "Flexible Dilatometer" technique (Ladanyi, B., Coordinator, 1987). During the last several years, the U.S. Bureau of Mines has developed techniques for the geomechanical characterization of in situ rock masses using hydraulic borehole pressure cells (Lu, 1981, 1986b, 1987). Lu (1988) proposed that comprehensive geomechanical characterization of a rock mass can be made using a single installation of a combination of three USBM hydraulic borehole pressure cells installed in one borehole (Fig. 1).

ABSTRACT: In situ stress measurements made between 1200 and 1500 meters depth at Minnova Inc.'s Ansil deposit resulted in highly unusual in situ stress tensors seemingly unrelated to the normal regional field stresses observed in the Canadian Shield. Further analysis of the field and numerical model data suggests that the massive sulphide orebody is behaving as a soft inclusion relative to the brittle host formations. Finite element analysis was used to obtain a preliminary calibration to the measured field stress values and thereby to induce the most probable upper bound for the far field stresses. 1 INTRODUCTION An in situ stress measurement program was conducted during the winter of 1988 at the Ansil mine, located 500 km northwest of Montreal, Quebec, Canada. The massive sulphide deposit is located between 1200 and 1500 meters below surface. It has an ellipsoidal mushroom shape varying in length from 100 to 140 meters in the upper Zone. The flattened tail of this "mushroom" is 15 meters thick and has a length of 365 meters, dipping at 50 degrees east. A peculiarity of the orebody is the almost 90 degree rotation between the strike of the upper and lower zones. The core of the orebody is rich in chalcopyrite (CuFeS2) and pyrrhotite (FeS) while the outside shell contains mostly sphalerite (ZnS) and pyrite (FeS2). Initial far field stress assumptions were based on published regional data, (Herget, 1982, 1987). The vertical stress was assumed to increase linearly with depth based on an estimated average density of 2700 kg/cu.m. The average horizontal stress was considered to be extremely high based on measurements made elsewhere in the region. Herget's relationship (Figure 1) results in a horizontal to vertical stress ratio of 2.6 at a depth of 1200 meters, (σv = 32 MPa; σ HMAX = 84 MPa). 2 HISTORY OF THE ANSIL DEPOSIT The Ansil copper deposit was discovered in 1981. A 1500 meter deep circular production shaft was sunk in 1984 and drill drifts were driven on the 7, 8 and 9 levels for final orebody delineation. The circular concrete lined production shaft is 6.71 meters (22 ft) 0.D. and 5.49 meters (18 ft) I.D. The total cost for developing the project to the production stage was $65 million Can. The ore zone is situated between a footwall rhyolite and a hangingwall andesite. Each formation is locally chlorite altered adjacent to the orebody and contains roughly three joint sets widely spaced. Index testing of the key rock units results in the average elastic properties listed in Table 1. Drill core RQD's suggest that the massive sulphide can vary from very good, (RQD = 100%), to very poor, (RQD = 20%). As more exposure of the ore has become available it has been observed that, although it is only lightly jointed, (one horizontal joint set), the material itself is often very weak and can be broken by hand. The sulphide was observed to break along small pre-existing fractures that often appeared rusty.

ABSTRACT: Backfill, particularly hydraulically placed, uncemented, deslimed tailings, is being used increasingly for support in the deep level stopes of the South African gold mines. Non-linear stress analyses are described, the results of which show good agreement with observed behaviour. The results of multi-step mining and backfilling analyses are compared with those of single step analyses, and it is concluded that the latter are satisfactory, particularly for comparative situations. Finally, mining at a depth of 5000m, using soft backfill for support, is analysed, and it is concluded that such support will provide adequate support at great depth. 1 INTRODUCTION The depth of operation of gold mines in South Africa is continually increasing and, at present, the deepest mining production is taking place at some 3400m below surface. Mining at such depths is a very difficult operation - travelling distances are great, transport facilities are limited and transportation of men and materials is time consuming. In addition, temperatures and stresses are very high. Backfill was first used for support in the South African gold mines eighty years ago. Its use was subsequently discontinued, however, and it is only in recent years that it has been reintroduced. Some of the delay in this reintroduction is no doubt due to the results of elastic mining simulation stress analyses, which indicated that significant benefit could only be gained from the use of backfill for support when spans were such that significant closure occurred in the stopes (Ryder and Wagner 1978). However, results of trial filling programmes showed that even soft, hydraulically- placed deslimed tailings fill provided good local support, and appears to be effective under seismic loading. In addition, it allows easier control of ventilation, an important aspect in the very hot mining environment. The long-recognized benefit of improved regional support is also present (Piper and Ryder 1988, Ryder and Wagner 1978). In this paper reported cases illustrating the effective behaviour of hydraulically placed backfill will be summarised briefly. A successful method of analysis for the backfill-rock mass interaction will then be dealt with, and finally the application of this method to the prediction of conditions resulting from mining at a depth of 5000m will be described. 2 REPORTED EFFECTIVENESS OF BACKFILL Substantiation, by measurement, of the effectiveness of backfill is limited at this stage. There have been several reports, based on observations, of the performance. Gay et al (1988) present two cases in which rockburst effects were recorded. In the one case, gullies in filled panels remained in good condition. Damage was minimal where the fill was close to the face. Fill was being placed at the time of the burst, and the fill paddock was slightly damaged, but no liquefaction of the fill occurred. In the second case conditions remained good in totally filled panels and there was only limited closure. Gulley conditions remained good. In partially filled panels substantial closure occurred.

ABSTRACT: In the first part, it is explained how the notion of scale effect, the data acquisition and statistic analysis as well as the characterization of the morphology of a fracture under stress, the assessment of interactions between structure, stress and fluid flow and the knowledge of the joint filling can be employed to improve our understanding of rock discontinuities. At the beginning of the second part, the recent developments of three-dimensional modelling of a single and a fractured medium are presented. Then, it is underlined that the special distribution of discontinuities can be deduced from geophysical surveys. INDUSTRIES INVOLVED IN RESEARCH ON ROCK DISCONTINUITIES Predicting the role played by rock discontinuities in fluid migration and stability of structures at great depth is an old challenge presented to scientists by many industrial organizations. These include firms involved in mining and civil engineering as well as those in the production of geothermal energy in nuclear waste storage and oil recovery which have sponsored the research presented at this conference. 2 TOWARDS A BETTER KNOWLEDGE OF THE BEHAVIOUR OF ROCK 2.1 Scale effects In order to understand the behaviour of a discontinuous rock mass, Hoek and Brown (1980) introduced the notion of scale effects. So, a fractured rock with a spacing relatively large compared to the size of the project can be analyzed using a continuum approach, while the modelling of a continuous medium with a single discontinuity in the zone concerned requires the use of the displacement discontinuity method for example. A highly fractured rock can be treated as a continuum with equivalent properties, but in other cases the block interactions must be taken into account by using an approach based on the distinct element method or block theory. When a borehole collapses, thermal or poro-elastic effects induced by the mud circulation. Nevertheless, it is important to consider the scale effects presented above and to distinguish the major discontinuities from the minor ones in order to minimise the risks of borehole collapse. As Maury (1989) underlined in his case history, it is only the block structure of the rock mass that can explain the loss of wellbore Pont d'As 5 drilled at 4500 m depth below Pau. The re-pressurisation of a fault triggered small shear movements, thus weakening the casing so that it could not withstand a decrease of internal pressure. Joints or planes of weakness with about ten centimetres spacing can also lead to instabilities at a borehole wall. Goodman (1976) analyzed the effects of joints by inserting their stiffness and spacing into the elastic matrix of an anisotropic material; Amadei (1983) proposed a general analytical solution for the stress distribution induced by the drilling of a borehole in such a medium subjected to a homogeneous tectonic in-situ stress field and induced by the pressurisation of the borehole. By coupling Goodman's and Amadei 's solution to failure criteria for rocks, Aadnoy (1988) showed that borehole instability can result from certain choices of the borehole inclination and azimuth.

ABSTRACT: When constructing underground openings at great depths, particular problems resulting from the squeezing characteristics of rock masses are often encountered. Large plastic deformation is one of the most critical problems and it must be considered in design practices. In order to stabilize rock masses, support structures such as steel ribs and rock bolts are commonly installed. However, no reliable design method is presently available for dimensioning these support structures, particularly rock bolts. This may be mainly due to the fact that the function mechanism of the bolts has not yet been fully investigated. In this paper, we shall discuss the effect of rock bolts based on field measurement and laboratory experiment data, and propose an estimation method which is capable of evaluating the effect of rock bolts. 1. INTRODUCTION When constructing underground openings at great depths, engineers are confronted with particular problems resulting from the squeezing characteristics of rock masses. In the past, the squeezing phenomenon has caused severe damage to several tunnels in Japan, and the repair work has involved a large consumption of money and time. Squeezing is a fundamental problem encountered when assessing the stability of underground openings at great depths. Its complexity arises from the inter-relationship of in-situ stress, material properties, and the rock supporting system. The first two variables are totally governed by natural conditions, and the third (the rock supporting system) is the only one over which humans can exert control. We hereby define 'some control' as referring to the stabilization of rock masses. Rock bolts have been recognized as one of the most effective supporting measures for stabilizing rock masses around underground openings. As far as the authors are concerned, however, there seems to be no reliable design method presently available for dimensioning a rock bolt support system. This may be mainly due to the fact that the function mechanism of rock bolts installed in rock masses has not yet been fully investigated. In practice, therefore, the design of rock bolts is entirely based on the experience of and judgement by well trained engineers. Engineers always run into difficulties, however, and face questions of how to determine the number and length of the rock bolts necessary to stabilize the rock masses. The purpose of this paper is to give answers to the above questions by proposing a method which is capable of quantitatively evaluating the effectiveness of rock bolts. The effectiveness can be assessed by comparing the strain occurring in the surrounding rock masses with the allowable strain. 2. THE STABILITY OF ROCK MASSES AT GREAT DEPTHS Many investigators have observed that the vertical initial stress component increases linearly with depth, and that the increase is related to the overburden pressure (see Fig. 1). On the other hand, the strength of rock masses does not drastically increase according to depth. This means that the large deformation/failure of rock masses becomes a serious problem, and should be considered in the construction of underground openings at great depths.

ABSTRACT: The tunnel at depth is stressed by the higher stress levels induced by excavating and is easily failed. Based on the deformation and failure behaviours of rock, it is investigated that the progressive failure of the rock mass around a tunnel at depth. And the analytic formulas are given to evaluate the effects of the softening and dilation of the rock mass on the stability of underground tunnel. Some suggestions are also given to improve the support of underground tunnel at great depth. 1. INTRODUCTION The stability of tunnel depends more on the response of the rock mass to the higher stress levels induced by excavating or mining activities. When a tunnel is made at great depth, the redistribution of stress usually results in stress levels high enough to cause progressive failure of the rock mass around tunnel, that is, the stability of tunnel will be lost slowly following excavating. Previous in-situ investigations of the tunnels at great depth have shown the existence and extent of the broken zone around such tunnel. It is necessary to take consideration the progressive failure behavior of rock mass in the design and construction of the tunnel at depth. Unfortunately, only a few of people take this idea for underground tunnel. The design and construction of underground tunnel are according to human experiences or classical theoretical results. An important feature in the design of underground tunnel is to utilise the rock itself as the principal structural support material by minimizing the disturbance of the rock mass during the excavation process and thus employing the minimum amount of artificial support such as steel arches, rings, concrete lining and shotcrete bolts. This is largely dependent upon geological conditions and the rock mass properties encounted in the tunnel. The lack of knowledge concerning the fundamental of rock mass failure has led to the development of numerous emperical and theoretical design formulae (Yuan &Chen 1987, 1988, Wilson 1980, Brown 1983 and Dritz 1984). Such design criteria are theoretical or conceptional and although they have proved to provide successful guidelines in particular circumstances, they currently lack precison and consistency in the variable rock mass and stress enviromments. A clear understanding of the behaviour of broken rock as well as the induced stress enviromment is necessary yet to further advance the effectiveness of such formulae for predictive purpose. The development of design method for underground tunnel support must include the criteria which consider the strength and deformation characteristics of rock mass. Knowledge of the complete progressive failure behaviors of rock mass around tunnel is therefore a prerequisite to the effective design of tunnel at increased depth, particularly in weak ground. A elastic-plastic medium has been made to model the progressive failure of rock mass. The strain softening and linized dilation model arc used to analyse the stress in rock mass around tunnel. An analytic formula is developed to evaluate the effects softening and dilation of the rock mass on the stability of underground tunnel.

ABSTRACT: A two dimensional computer model based on the theory of stochastic process is applied to simulate time-dependent failure and deformation of rock mass around underground roadway in deep mine. In this model, the failure process of rock is assumed as a probabilistic rate process in which the rate constant is given by a power function of stress severity, and FEM technique is used to model deformation of rock mass and rock bolt supports. According to which of shear or tensile is the failure mode, different anisotropic mechanical properties are assumed for post-failured elements in the fractured zone. Through the computer modelling, it is found that grouted rock bolt can retard the convergence of roadway, however, at great depth, the effect of rock bolt is not so remarkable as expected from experiences at present mine roadways. 1 INTRODUCTION Strata control is one of the most important problems in deep underground mines, where many roadways suffer from severe rock pressure. The severity of rock pressure depends upon the depth of roadway as well as the strength of rock formation around the roadway. In other words, it depends upon the competence factor which is defined as the ratio of uniaxial compressive strength of rock to the overburden pressure. In Japan, most of coal bearing formations belong to Tertiary system, and coal measure rocks are frequently weak and fractured. Most of output are mined from coal faces at the depth of several hundreds meters from ground surface. Furthermore, various non-ferrous metallic ore deposits are frequently buried in weak volcanic rock formations (Nishimatsu, 1981). Frequently, the rock formation is assumed as an elastic body for theoretical analysis in rock mechanics. However, it is indispensable to take time-dependent behaviours of rock into account for analyzing or modelling the convergence of mine road way, because the elastic deformation takes place instantaneously after face advancing or heading, and gradual time-dependent increase of convergence and support load can not be explained by elastic theory. Furthermore, progressive failure or yield is frequently observed around underground openings. The progressive failure would play an important role in the time dependent rock pressure phenomena, and be assumed as a rate process to consider effect of time elapsed. For examples, creep of rock sample has been analyzed under the assumption that crack extension is a probabilistic rate process (stochastic process)(Scholz, 1968). Recently, the authors have assumed that crack growth is a stochastic process and conducted a computer modelling of time-dependent deformation and failure process of rock samples under various test conditions (Okubo and Nishimatsu,1986). In this paper, under the same assumption as the previous paper, the authors conduct a computer modelling of sometime-dependent behaviours of rock formation around underground mine roadway by means of F.E.M.technique. 2 COMPUTER MODELLING 2.1 Basic assumptions For the purpose of modelling the time dependent behaviour of rock, it is assumed that time-dependent component of strain appears as a consequence of reduction in the effective elastic stiffness of rock resulting from localized brittle fracture or crack extension.

ABSTRACT: In this paper, the mechanism of tectonic earthquake is discussed qualitatively, and the sequence of Tangshan earthquake is simulated with aid of computer. On the basis of this, a group of mechanical parameters and some qualitative understanding for the deeply buried rocks is gained. 1 INTRODUCTION Since the deeply buried rock is the direct object of earthquake occurrence, it is certainly significant to get a clear understanding of mechanical behavior of deeply buried rocks for the study on the mechanism of earthquake forming, generating and the long term prediction of earthquakes. But as things are, although researches have been done a great deal of rock experiments; and obtained mechanical parameters under various forcing condition rock specimens used in experiment were mostly from the place where were easy to sample in the upper crust. Obviously, it is not proper to apply these experimental results in deeply buried rocks. We now generally infer that the deeply buried rock is in the state of high-temperature, high-pressure and high-density. Many researchers have done experiment under similar conditions (Stesky, 1975; Byerlge, 1975; Engelder, 1975). But how to apply their results into the actual situation is a problem pending further study in itself. However, from their studies, we gained a good deal of enlightenment and experimental basis. It is well known that the earthquake carry a great number of information, we may take it as an information source, and try to re-appear the forming and generating procedure of occured earthquakes with the theory of seismology and morden mechanics. By this, quantitatively and qualitatively, we will gain some mechanical parameters and properties of deeply buried rocks. 2 THE MECHANISM OF THE TECTONIC EARTHQUAKE ORIGIN It is well known that the earth is always in moving, consequently, there is force in the crust. When the force, whose magnitude and orientation are different with the space and the time, act on the rock layer which makes up the crust, the crust certainly will deform, such as fracture, fold, ascent or descent of land and large scale magmatism, meanwhile, earthquake is caused. This kind of earthquakes is called tectonic earthquake, it is the most common in the crust, and the majority of shallow focus earthquake (h≤ 70km) belong to this kind, which are often destructive shocks. The tectonic earthquake usually occur at or near the fault terminus or at the locking area. Because of the inhomogeneity of rocks, the displacement and stress field are always in adjusting, it result that around the rocking area or the fault turminus stress will concentrate to a high degree, and a great deal of elastic energy will be accumulated. When the stress level exceeds the locking strength or the frictional stress, the dislocation along the fault plane will be caused, the accumulated strain energy will be suddenly released and the static frictional stress on the dislocation surface may drop considerably at the time of rupture because of the loss of cohesion.