Rapid development of world economy has been gradually exhausting the shallow mineral resources. More and more attentions have been paid to the mineral exploitation in deeper subsurface of the earth. For examples，current coal mining has reached 1 500 m in depth; geothermal exploitation has reached over 3 000 m，the depth for ferrous metal mining over 4 350 m and for oil and gas development 7 500 m. Therefore，deep mining is an on-going mining industry. However，more and more engineering accidents or hazards have occurred in deep mining. These accidents or hazards are difficult to be predicted with the current theories or experiences because the mechanical behaviors in deeper ground have not been well understood. On the other hand，current research and development in rock mechanics cannot solve the practical problems in deep ground engineering. A novel research and development scheme should be specially designed for deep ground engineering. Particularly，fundamental concepts and basic theory of rock mechanics should be revisited for deep ground engineering. For examples，what is the deep ground engineering？Is it measurable by the depth of ground？What is the essential difference of the mechanical properties of rock in shallower and deeper subsurface？Can the classical rock mechanics be applied to describe the mechanical behaviors of rocks in deep ground？How can the black box of rock mechanics in deep ground be revealed during the mining induced disturbance？How does engineering activity such as the development and storage of energy，CO2 sequestration and nuclear waste disposal in deep environments(such as earthquake，geology，geochemistry，geothermal environment as well as microbial environment) affect the micro-scale variation of rock？These are the fundamental questions in rock mechanics for deep ground engineering. This paper revisits these fundamental concepts by taking the rock mechanics in coal mining as an example. Our results showed that hydrostatic pressure was a typical stress state in deep ground engineering. Deep ground was defined by its stress state instead of its depth. Based on this concept of stress state，we proposed a framework of rock mechanics for mining process in terms of stress paths. Such a framework is different from the classical rock mechanics. Furthermore，a visualization technology was then developed for the investigation of rock mechanics in deep ground. This technology combined the CT scanning，the 3D printing，the reconstruction technology with fractal analysis and the stress freezing technology. This visualization technology provided transparent observation of the disturbance of stress during mining，the development of fractures，the volumetric breaking，the plastic instability and the micro-seepage.

The elongation and impacting experimental system was used to study the dynamic mechanical properties of the constant resistance and large deformation(CRLD) bolts，such as the impacting force，the expansion and elongation of one type of CRLDs. The negative Poisson?s ratio effect was verified and the experimental results were compared with the ones from finite element analysis(FEA). The experimental results showed that the bolt absorbed the impact energy through a process of elastic-plastic structural deformation with the unique negative Poisson?s ratio effect and was well adapted to the dynamic impacting process. The FEA results were matched with the experiment ones，which not only confirmed the accuracy of the FEA method，but also verified the feasibility of using FEA method to prove or substitute the experiment.

An overview of the investigations on coal bumps in China is presented from three aspects：mechanism，forecast and control. The key issues and the corresponding scientific achievements of investigations on coal bumps in recent years are reported. These issues involve the geological conditions related to coal bumps and the quantitative analysis method，the time-space distribution of mining-induced stress and energy in mining disturbed region，and the spatial structures of overlying strata and its mechanism of coal bump triggering. Several technical systems of monitoring and warning for coal bumps are introduced. These systems involve the integrated monitoring system of stress and seismic fields，the monitoring and detecting system of coal bumps，the ZOS distributed micro-seismic monitoring system，the wireless monitoring system of coal bumps and the electrical charge monitoring technique. Finally，some new techniques for controlling coal bumps and related application in mining groups of Datong and Yima are discussed. The control methods include the energy-absorption and anti-scour support，the anchor bolt with constant resistance and large deformation，and the regional prevention technical system based on stress control.

A joint factor reflecting the distribution characteristics of columnar joints was introduced to represent the influences of joints on the strength of columnar jointed rock masses based on the nonlinear strength criterion proposed by Ramamurthy. An anisotropic strength criterion was suggested for the regular hexagonal columnar jointed rock masses with the strength nonlinearity being described with the power function. The method for calculating the joint factors of regular hexagonal columnar jointed rock masses was investigated，and the constants in the anisotropic strength criterion were determined with the uniaxial and triaxial compression results from tests simulating the columnar jointed rock masses. The experimental results on the strength of jointed rock masses published in the past literatures were used to validate the proposed strength criterion preliminarily. From the validation，it was found that the predictions from the proposed strength criterion agreed well with the experimental data，indicating that the anisotropic strength criterion described well the strength anisotropy and nonlinearity of columnar jointed rock masses.

Without timely and reasonable treatments，the large convergence displacement due to high geostress and complex tectonic stress of ground squeezing will result in economic losses. TBM tunnels were the focus of the study. The mechanism of large displacement due to ground squeezing was described. The prediction and recognition methods were presented. The prediction methods of convergence displacement of tunnel including AI method and uncertainty analysis method were discussed in detail. Finally the common disposal methods for grounds under different squeezing levels were summarized. The formulas for recognition，the prediction methods and the disposal technologies of large squeezing displacement of TBM tunnels were developed for corresponding projects，so blindly using these methods yielded no effect. We suggested the choosing of these methods with caution.

To solve the supporting issues related to the roadways at depths more than 1 000 m in Xinwen coal mining area，the mechanical parameters，such as the in-situ stresses，the strengths and the structures of surrounding rocks，the deformation and damage states of surrounding rocks of deep roadway and the supports in coal mines， were analyzed. The deformation and damage characteristics of rocks surrounding a rock roadway with depth more than 1 000 m under various supporting types and parameters were simulated with UDEC. Based on the in-situ test data and numerical simulation results，the full section reinforcement with the high strength and pretension rock bolts and cables combined with grouting was put forward for the support of No.1180 main return roadway in Huafeng coal mine，Xinwen coal mining area. The underground test was described in detail，including the support parameters，materials，underground construction technologies for cables and grouting in roadway floor，and monitoring results. The supporting effects were evaluated through the analyses on the data of surrounding rock displacements，roof separations and the loads along bolts and cables. The underground test results confirmed that the large deformation of surrounding rock in the main return roadway with depth more than 1 000 m was effectively controlled，and the long term stability was kept by means of the combined support system of high pretension and strength bolts and cables with grouting.

The total length of Jinping II deep buried tunnel cross different geological units is more than 16 km. In the process of excavation，rockburst was controlled by the geological structure conditions and the local stress field. The tectonic rockburst is directly affected by the geological tectonics. Based on the systematic analysis of the forefront research results about rockburst in deep engineering home and abroad，and combined with the rockburst revealed by a large number of cases in Jinping II deep buried tunnel engineering，the tectonic rockbursts were further subdivided into end tectonic，slip tectonic and strain tectonic. The inverse analysis of the rockburst was carried out with the discrete element program and related risk evaluation index of rockburst. The mechanism of different kinds of tectonic rockburst in deep buried tunnel engineering was revealed，and the key controlling factors of rockburst and control mechanisms of various factors were analyzed. Finally，the corresponding control schemes were put forward according to the different types of tectonic rockburst.

To seek a simple and effective method for the calculation of the stress release ratio in tunnel excavation，a revised method was proposed via the volume loss ratio based on the principle that the stress release ratio was approximately equivalent to the release ratio of displacement. The calculation of the stress release ratios for tunnel in soft rock under different construction schemes was performed in detail and thus the variations of stress release ratios under different construction schemes were analyzed. The calculated results indicated that the stress release ratios were different for various excavation parts of different schemes. In general，the stress release ratio of latter excavated was larger than that of the part excavated earlier due to the disturbance effect of excavation. The average stress release ratio decreased gradually with the improvement of the schemes except for the CD excavation method.

Valley deformation induced by impoundment affects the current working behavior and long-term safety of high arch dams. However，the evaluation criterion for the influence of valley deformation is still limited in the design method of arch dam. In this paper，the features of valley deformation during impoundment in the site of high arch dam and the limitation of conventional calculation methods were summarized. Based on the effects of water pressure in fractures，the hydrostatic pressure was taken into consideration in the yield function of elastoplastic model，and applied in the numerical simulation of Jinping I arch dam. The calculated displacements agreed well with the observed values，which preliminarily explained the reduction of valley width and other deformation phenomena in the dam site. The mesoscopic deformation mechanism of fractured rock mass during initial impoundment was then explored. The effective stress principle for unsaturated and non-persistent fractured rock mass was presented. The research revealed that the water pressure distributed in fractures changed the equilibrium state of rock mass，which was the main driving force of plastic deformation in the dam site after impoundment.

Variation of wave velocity with stress in rock is very significant for the study of damage development. Nine different models were set with the particle flow code(PFC) to simulate the processes of wave propagation and attenuation，in which a velocity pulse was applied to the transmitter and the average velocity was recorded with the receiver. It was revealed that the coordination number was in linear relationship with the porosity. The branch vector distribution was directly related to the amplitude attenuation. Moreover，the wave velocity was mostly affected by the porosity，coordination number tensor and stiffness tensor. The stress-strain process was simulated under biaxial compression to explore the wave propagation. The micro cracks were mainly distributed along the axial direction. In comparison with the shear cracks，the angles between the dominant orientation of the tensile cracks and axis are smaller. With the increase of micro cracks，the anisotropic degree of wave velocity and the distribution of branch vector increased gradually. It was proved that the intrinsic reasons of the velocity variation were the formation of new contacts，the breakage of bonds and the separation of contacts. The wave velocity was essentially consistent with the square root of the corresponding component of stiffness tensor. It was a new method to quantify the damage and wave velocity through analyzing the stiffness tensor and branch vector distribution，which provided reference for the study of damage development.

In order to determine the reasonable bolt length for Jinping II deep buried tunnels，the statistical analysis to the failure mode of surrounding rock was carried out and the main problem in bolt design was identified. Because the crack development in the damaged zone led to the decrease of wave velocity，the distribution characteristics of the damaged zone were obtained through the summarization of the typical acoustic test results in the tunnels. The related parameters of particle flow code(PFC) were fitted with the field acoustic test results，and the buried depths of damaged zone at different depths were forecasted. The results showed that the depth of rock damage and the number of cracks were also increased slowly with the increase of buried depth. This feature was beneficial to the design of the bolt length. The calculated results with PFC showed the phenomena of localized damage and that the surrounding rock in the damage zone still had the bearing capacity. The bolt installation in the damage zone still had the effect of limiting the damage development，and met the requirements of the damage    zone supporting.

Stability analysis of hard rock pillars is significant for hazards prediction to the mining of thick coal seams of extremely steep. Modeling the hard and extremely steep rock pillar experimentally was accomplished with the composite-loading facility. The temperature and acoustic emission(AE) parameters were recorded with the infrared thermal instruments and AE sensors. Areas of heat radiation on the surface of rock pillar were found to emerge with a low temperature area near the boundary，a middle temperature area between boundaries，and a high temperature area across the boundary. The hard rock pillars experienced different stages including the elastic deformation，the micro ruptures and the fracturing instability. The accumulation ratio of AE energy was increased continuously and the temperature in the radiation area was decreased gradually. The reasons of AE and temperature variation during the hard rock pillar failure were the thermo-elastic transformation effect and thermo-friction effect.

为了研究起爆时差对孔间裂纹扩展的影响，采用动态焦散线实验的方法，研究含缺陷介质相邻切槽炮孔被采用同时起爆和微差起爆时孔间爆生裂纹的扩展和贯通机制。研究结果表明：(1) 在定向断裂控制爆破过程中，其他参数相同时，同时起爆的爆生裂纹成壁质量比微差起爆的质量好。因此，在岩体的定向开挖工程中，可采用同时起爆的方法来提高定向断裂控制爆破的效果。(2) 爆炸应力波中的P波前缘压缩波与爆生运动裂纹相互作用降低了爆生运动裂纹的扩展速度和爆生裂纹尖端应力强度因子的值。实验研究结果对于定向断裂控制爆破工程实践具有一定的借鉴意义和参考价值。

To achieve the adjustability of the parameters of geostress，gas pressure and coal properties，some design ideas and technical requirements of the simulation system were proposed based on the hypothesis of combined effects and CSIRO model after analyzing the advantages and disadvantages of existing devices. The simulation system was developed with the modular design idea. The system can apply geostress up to 30 MPa and gas stress up to 3 MPa. The acquisition frequency of gas pressure is up to 1 000 Hz. Synchronized acquisition of gas pressure and speed achieved precise record of the prominent transient information. The system is easy to operate and the single test is fast which shorten the testing cycles. 9 groups of crosscut exposing coal outburst simulation test of gas filled coal were carried out. The test results showed that coal and gas outburst was affected by the combined effect of geostress，gas pressure and coal strength. The coal strength prevented the outburst. The dynamic thresholds of gas pressure existed corresponding to different damage states of coal. The dynamic phenomena appeared strongly，and the cavities were large inside with small exit holes after outburst.

Cracks in rock will initiate，propagate and coalesce under loading，which remarkably influences the strength and damage property of rock. Based on a DIC system which was proved reliable，uniaxial compression experiments were carried out on the specimens with two precast flaws. The strain field and strain localization of the whole specimen were obtained. It was found that the strain localization and the progressive damage were very distinctive in the test. The strain localization zone appeared in the initial phase governed the range and trend of the whole process of the strain field development. There were four failure modes found in the test，that is the non-coalescence mode，the coalescence mode at inner tip of flaws，the coalescence mode at outer tip of fracture and the inner-outer tip coalescence mode. The bridge angle and the arrangements of flaws had the effect on the formation of strain localization zone in mesoscopic level and determined the cracking path and the failure mode in macroscopic level. The coalescence of strain localization zone in the bridge area resulted in a fluctuation of stress-strain curve before the peak and reduced the peak strength. The slower the coalescence of the strain localization zones，the higher the peak strength.

Laboratory modeling tests were performed to study the elastic wave propagation and attenuation in deep-buried rock mass with joints and the velocity and amplitude variation of elastic wave in jointed rock mass under different in-situ stresses were obtained. The propagation and attenuation of elastic wave in jointed rock mass were analyzed to study the relationships of the velocity increment and attenuation of elastic waves，the confining pressure，the number and angles of joints. The results of a triaxial compression test with constant confining pressure and increasing axial compression pressure showed that the change of elastic wave velocity had four phases：rapid growth，slow growth，stable state and rapid reduction. When the axial stress reached 60% of its strength，the velocity reduced rapidly，which indicated that the rock started to expand. The attenuation of the first wave amplitudes showed the similar trend of variation. In a triaxial tension-compression test with constant confining pressure and increasing axial tension pressure，the elastic wave velocity reduced gradually while the attenuation of amplitude grew. The results showed that the confining pressure，the angles and number of joints were all influential to the velocity and the attenuation of elastic wave. The degrees of influence sensitivity were in the decreasing order as：the number of joints＞confining pressure＞angles of joints.

Prevention and curing of delayed water bursting in rock fault are extremely difficult. Multiple field information monitoring and warning on delayed water bursting in deep rock fault in Wanglou coal mine were carried out. The source and conducting channel of ground water were analyzed. The delayed water bursting in fault during the mining process was divided into different stages. The temporal and spatial ranges of monitoring and warning were determined and the monitoring thresholds of temperature field and seepage pressure field were divided according to the theory of key water resisting layer and geophysical prospecting results，so that the identification criteria for monitoring and warning were established. Accurate and effective assembling of sensors was fulfilled by connecting them in series into monitoring units and plugging the drilling holes into sections using the returning slurry technique. Furthermore，the online monitoring on surface was fulfilled by taking advantage of the optical fibre monitoring system. Results showed that the extension of fissures in fault was effectively controlled by the water resisting coal wedge，which reduced the transmissibility of fractured fault zone.

To study the coal roadway support，the coal roadway in Fenxi mining area was used as the prototype for similar model experiments. The similar materials of coal rock layer were determined by the matching experiment. The simulated structure forms of anchor bolts and anchor cables were designed and the implementation process was expounded. Three groups of models were established to simulate the coal roadway under the conventional support，the support with the sidewall strengthened and the support with both the sidewall and corner strengthened. The experimental data were analyzed，which verified the effectiveness of sidewall and corner strengthened support which was successfully applied in engineering practice. The results of the study showed that the sidewall and corner strengthened support increased the vertical stress in the roof and floor，reduced the vertical stress in the stress concentration zone in sidewall，improved the stress state of surrounding rock，thus formed the benign effect between roof，floor and sidewall. The technology reduced the absolute and relative displacements of surrounding rock. The relative displacement in the sidewall was larger than that in the roof of coal roadway，indicating that the integrity of sidewall was worse than the roof and tend to be locally unstable. The technology controlled the generation and propagation of the cracks，prevented the failure of the sidewall and roof and thus  improved the safety. The engineering application and monitored results in Xinyu mine showed that the technology had good adaptability.

The instable failure of laminated plate with low tensile strength was explained with the theory of elastic stability. The granite plates of 80 mm in width，270–420 mm in length and 8–20 mm in thickness were compressed using the rigid loading heads with or without the soft rubber cushions of 5–10 mm in thickness. The plates with thickness over 16 mm failed in laminated style and the strength did not increase with the increase of the thickness and was about 40% of the uniaxial compressive strength of standard cylinder sample. The thinner plates were described with Euler?s instability of compression. Similarly，the uniaxial compressive strength was much lower than that from the trend of strength with confining pressure due to axial instability of laminated rock. The surrounding rock with low tensile strength may be structurally instable and zonal disintegrated in deep tunnel under the axial stress.

A tensorial measure，called the crack fabric，defines the geometric distribution of cracks in rock. Based on the theoretical framework for fabric description with the principle of stereology，a quantitative method describing the crack fabric in rock was proposed. The crack tensor was redefined with the normalized quantities. The trace of the redefined tensor is a constant，which brings a great convenience for determining the crack fabric. The plane crack tensor of rock was defined with the test line. The amplitude parameters and the component angle defined with the second invariant and third invariant of the plane crack tensor respectively，which described the degree and the direction of crack anisotropy. The plane distribution of cracks was described with two defined scalar quantity. The three-dimensional orthotropic crack tensors were deduced uniquely with any two amplitude parameters of three orthogonal planes. With the geometry changes of the crack distribution，the orthotropic fabric tensor degenerated naturally into the form of transversely isotropic and isotropic. Verification of coal CT test results showed that the tensor described better the plane distribution of cracks.

In the tunneling of the Jinan rail transit line R1 passing under Lashan River，the stability of surrounding rock，the effect of shield thrust and the sensitivity of interface parameters were analyzed employing both the theoretical calculation and numerical method. The study showed that，as the shield passed through the medium interface，the settlement of surrounding rock at arch crown increased firstly and then decreased，while the principal stress in the lining segment increased firstly，then declined and rebounded. The interface effect reached 4 meters to the front and 10 meters in behind. If the safety factor was 2，the support stress ratios maintaining the stability of rock medium，soil medium and excavation face of watercourse were 0.24–2.80，0.34–1.65，0.54–1.40，respectively. The rock-soil interfacial effect was found to increase with the increase of stiffness ratio of medium and interface dipping angle. Especially in the situation of  and ，the disturbance sensitivity of interface surrounding rock thrust became more significant.