Instability Of Surrounding Rock Research Articles

Integrated early warning and reinforcement support system for soft rock tunnels: A novel approach utilizing catastrophe theory and energy transfer laws

The initial support strength of deep-buried soft rock tunnels during excavation is often inadequate, and the timing of reinforcement support is inappropriate. This inadequacy leads to significant deformations in the tunnel, causing the energy released by the surrounding rock to exceed the energy that the support system can absorb, ultimately resulting in tunnel collapse. Tunnel deformation monitoring and anchorage support, when combined with energy transfer laws, are thus imperative for studying collapse prediction, the timing of reinforcement support, and parameters of the reinforcement support system of the surrounding rock. In this paper, a novel approach for early warning and support control in tunnel construction facing large deformations is introduced, and a predictive model for surrounding rock instability utilizing catastrophe theory and energy transfer laws is developed. By analyzing the correlations of rock deformation and energy dissipation, the timing of anchor reinforcement support was determined using function fitting techniques. Further, an energy transfer model under reinforced anchoring support was established, and the energy absorption parameters of the support body were determined. Also, the support parameter design scheme was proposed. The results indicate that the early warning and reinforcement support scheme proposed in this paper reduces the release of potential energy by 35.71 %; decreases the number of steps needed to reduce kinetic energy to zero by 66.67 %; reduces the total dissipated energy in joint shear by 33.68 %; increases the total stored strain energy in the material by 5.70 %; and reduces the total dissipated work in plastic deformation by 26.15 %. Additionally, this plan effectively controls the originally predicted large deformation area of the meter level to within 350 mm.

Surrounding rock instability mechanism for fault-crossing tunnels in water-rich soft rock

Surrounding rock instability mechanism for fault-crossing tunnels in water-rich soft rock

Stress–Structural Failure of a 610 m Crushing Station Left-Side Tunnel Section in Jinchuan II Mine: A Numerical Simulation Study

To address the stress–structural failure phenomenon that can be induced by the excavation of a left-side tunnel section of a 610 m crushing station, an unmanned aerial vehicle was used in this study to collect the geological conditions and rock mass information of the working face, and important geometric information such as the attitude and spacing of rock mass were extracted. Based on the identified attitude and spacing information, a three-dimensional rock mass structure and numerical simulation model of the 610 m crushing station left-side tunnel section were constructed using discrete element numerical simulation software (3DEC) (version 5.0). The results show that the surrounding rock instability of the left-side tunnel section of the 610 m crushing station is controlled by both the stress field in the contact zone between reddish-brown granite stratum and the gray-black-gray gneiss stratum. The cause of stress–structural failure is that the joint sets (JSet #2 and JSet #3) are most likely to form unfavorable blocks with the excavation surface due to unloading triggered by the excavation. Therefore, stress–structural failure disasters in jointed strata sections are one of the key issues for surrounding rock stability during crushing station excavation. It is suggested to adopt ‘optimized excavation parameters + combined support forms’ to systematically control stress–structural failure after unloading due to the excavation from three levels: surface, shallow, and deep. The stress–structural failure mechanism of deep rock mass is generally applicable to a large extent, so the results of this research have reference value for engineering projects facing similar problems around the world.

Study on Instability Change Law of Surrounding Rock of Tunnel Crossing Fault Fracture Zone Based on Cusp Point Mutation Theory

The instability process of surrounding rock is a nonlinear mutation. To truly reflect the nonlinear mutation in the failure of surrounding rock and grasp the opportunity of instability of surrounding rock, it is necessary to study the stability of surrounding rock by combining the existing methods with nonlinear scientific theory. Based on the numerical simulation results of a tunnel project crossing the fault fracture zone, the mutation theory is introduced to study the instability law of surrounding rock during tunnel excavation. The dynamic excavation process of surrounding rock tunnel in fault fracture zone is simulated by the finite difference software FLAC3D. Based on the cusp point mutation theory, the deformation law of each excavation step to the working face under different dip angles, strike angles and thicknesses of fault fracture zone is studied. Combined with the dichotomy method, curve fitting is used to analyze the instability threshold of surrounding rock and predict the opportunity of surrounding rock instability. The results indicate that (1) As the dip angle and strike angle of the fault gradually increase, the settlement of the surrounding rock arch first decreases and then increases, while as the thickness of the fault increases, the vertical displacement of the surrounding rock gradually increases. (2) When the dip angle of the fault is less than 90°, the sudden instability of the surrounding rock is advanced, while when it is greater than 90°, the time for sudden instability is delayed. (3) As the strike angle and thickness of the fault gradually increase, the excavation step for the tunnel to undergo sudden instability is gradually advanced. The research conclusion can provide theoretical guidance and suggestions for similar engineering construction in the future.

Dynamic Response and Failure Mechanism of Deep-Buried Tunnel with Small Net Distance under Blasting Load

Under blasting load, a series of safety problems, such as lining cracking and surrounding rock instability, are prone to occur in deep-buried tunnels with a small net distance. It is significant to understand the dynamic response and failure mechanism of tunnels under blasting. The blasting attenuation formula is optimized through theoretical analysis and field experiments. The measuring point vibration is monitored in real time and the tunnel blasting model is established by ANSYS/LS-DYNA software. The model was set as having no reflective boundary and an uncoupled charge structure was used. The attenuation law of blasting seismic waves is studied from the adjacent tunnel lining and the direction of the tunnel cross-section and length. The inner and outer sides of the tunnel lining are investigated, respectively. The displacement and acceleration of lining measuring point are also analyzed. The dynamic response of the tunnel lining under blasting excavation is analyzed from multiple angles. The results show that the arch foot on the inner side of the lining (the side in contact with the tunnel headroom) is the first to generate vibration. On the outside of the lining (the side in contact with the rock),the peak vibration velocity is reached after blasting load unloading. There is little difference in the vibration velocity at different positions of the transverse section, but great difference in the vibration velocity of the longitudinal section. The influence of the horizontal displacement was greater than that of the vertical displacement. The vibration acceleration of the measuring point at the arch foot of the section is the largest and the detonation is also the largest.

Study on surrounding rock failure mechanism and rational coal pillar width of the gob‐side coal roadway under influence of intense dynamic pressure

AbstractA reasonable width of the coal pillar is very important to the surrounding rock control of the gob‐side roadway. An unreasonable width of the coal pillar will make the roadway to be located within the range of strong mining influence, leading to severe deformation of the roadway. Severe subsidence at the coal pillar side of the roof and serious coal pillar deformation are problems caused by strong dynamic pressure due to mining in the gob‐side coal roadway. This paper studies the surrounding rock instability mechanism and rational coal pillar width of the gob‐side coal roadway under the influence of intense dynamic pressure. The results show that: (1) Under the condition of large mining height, the roadway overburden is a hinged structure, and an unreasonable coal pillar width makes the gob‐side coal roadway to be located below the main roof fracture line. The rotary movement of the key block of the main roof is the main reason for the roadway deformation. (2) According to the evolution law of stress field, displacement field and plastic zone of surrounding rock of roadway under different coal pillar widths during roadway driving and mining were studied, and it is concluded that a 6‐m‐width coal pillar is the most reasonable. (3) Based on the stress distribution and plasticizing range of surrounding rock in a narrow pillar roadway, the combined support scheme of “anchor cable + grouting + single prop” is proposed and applied to engineering practice. The practice results show that the roadway deformation is well controlled, and the safe mining of the working face is realized.

Experimental study on precursor characteristics of rock failure based on strain and temperature changes

Early warning of natural disasters (earthquakes and landslides) and engineering disasters (mining and tunnel collapse) has always been a complex problem in the field of engineering. In the current paper, the uniaxial compression loading tests of rock specimens are methodically observed by employing strain data acquisition equipment and an infrared thermal imager. By examining the variation pattern of strain field and temperature field data before rock damage, a critical damage identification method in conjunction with the strain and temperature data analysis is proposed. The corresponding time of total failure under the action of uniaxial compression is rationally determined by the strain data variation from increasing to decreasing in a 45-degree direction, and the location of the crack is predicted based on the slope of the tangent line of the strain curve in x- and y-direction at the failure time. Additionally, the slowing down of the temperature growth trend at the monitoring point and the significant variation of the local temperature over a short period in the infrared thermal image could be utilized as precursory characteristics for the rock damage. These research findings could be very beneficial in predicting the early warning of the surrounding rock instability in geotechnical engineering.

Instability energy mechanism of super-large section crossing chambers in deep coal mines

The stress concentration and failure at chamber intersections in coal mine are intense, especially in deep-buried, super-large section conditions. In this paper, the plastic radius of super-large section chamber under unequal pressure was corrected on the basis of the size effect. Then, stress and failure evolution of intersections under different crossing angles and equivalent angular bisectors were revealed. Furthermore, 2 trajectory curves of failure and stress were analytically expressed, which divided the intersection into 5 influencing zones in the light of stress superposition degree. After determining instability trigger point and instability path, instability energy criterion of intersection can be obtained as K > 1, which means that the external energy is greater than the sum of energy consumed by surrounding rock instability and supporting structure failure. Taking coal-gangue separation system of Longgu Coal Mine as example, it was found that there was instability risk under original parameters. For long-term stability, an optimization design method was proposed by considering safety factor, and optimal support scheme was obtained. Field monitoring showed intersections deformations were relatively small with the maximum of 125 mm, which verified the rationality of theoretical analysis. This study provides guidance for the stability control of the intersections under the same or similar conditions.

An Improved Method for Estimating the Strength of Jointed Rock Mass Using Drilling Technology

During underground excavations for civil and mining engineering purposes, the variations in mechanical properties of jointed rock mass are important, especially for the design of underground structures. The unreasonable evaluation of strength properties for rock masses can lead to the structure instability of underground construct, which subsequently causes collapse accidents in underground engineering and poses a serious threat to the structures safety. Rock mass classification system is an effective method for determining the strength parameters for rock masses. Firstly, the existing approaches for determining the strength parameters of rock masses are reviewed using classification systems such as the joint factor, jointing index methods, rock mass rating (RMR), Q-system (Q), geological strength index (GSI), and rock quality designation (RQD). Since in some cases, rock quality designation (RQD) is the only information available, a comparative study with other methods then is conducted to verify the limitation and deficiencies of RQD method. On this basis, a newly developed method from RQD classification system based on drilling special energy is established to estimate the rock mass strength. Finally, the developed method is applied in the tunnel engineering in the Daheba hydropower station in Hanjiang to the Weihe River project of China, and the results are compared with those from the previous works. The in situ test results show that the developed method can give appropriate values compared to various suggested relationships from RMR, Q, and GSI. The results can be treated as an important design reference for the prevention and treatment of surrounding rock instability and underground excavations in the Daheba hydropower station.

Numerical Analysis on the Joint Weakening Effect of Rock Mass Behaviors in Tension

Abstract The presence of joints and other types of discontinuities has a significant effect on the mechanical properties of rock, especially for tensile properties to fundamentally influence the stability of rock excavations. The main challenge associated with the experimental research on jointed rock lies in the difficulty to carry out amount of direct tensile tests for analysis of the effect of joint geometric parameters on mechanical properties. In this study, a particle flow model was established by utilizing the flat-joint contact model (FJM) to represent the rock materials. After microscopic parameter calibration, 53 sets of the numerical model were used for investigating the relationship between jointed geometric parameters and tensile mechanical properties. The results show that the crack initiation is related to trace length l and joint angle β, and the tensile-shear crack will appear as β increase. The uniaxial tension strength σt and β had first a weak negative correlation and then a positive correlation as the β increases, which was consistent with mathematical calculations. Furthermore, the relative importance (RI) analysis showed that the β plays a decisive role among the joint geometric parameters for affecting σt, and the effect factors of σt were joint angle β, length l, density n, and aperture d in that order. The present research can be utilized for multiple purposes in the field of jointed rock engineering, such as prediction of surrounding rock instability analysis and estimating the variable values in the inversion analysis in practical engineering projects.

Quantitative Risk Analysis of Tunnel Instability in Layered Rock Mass Considering the Spatial Variability of Elastic Modulus

Layered rock mass is a typical excavation disturbance sensitive rock mass, which is prone to the large deformation and collapse of soft rock mass under tunnel unloading. This in turn causes the construction safety risk and delays the project construction. As the spatial variability of rock mechanics parameters existed widely, which could make the deterministic calculation results cannot really reflect the safety reserve of the rock mass by supports. Therefore, considering the spatial variability of elastic modulus of layered rock mass for study the risk of tunnel instability both has the theoretical and practical significance. According to the transverse anisotropy characteristics of layered rock mass in Baokang tunnel, China, elastic wave velocities were determined by the acoustic test method at different broken surrounding rock mass. Meanwhile, the elastic modulus was converted into the elastic modulus based on theoretical formula, and a database of the elastic modulus was set up. Using the variograms model of geostatistics, the autocorrelation distances of elastic modulus along the strike and dip of rock strata were determined, respectively. Furthermore, a two-dimensional parameter random field which can characterize the spatial variability of elastic modulus of layered rock mass was constructed by the Karhunen-Loève (K-L) expansion method. Taking the standard value of tunnel allowable deformation as the evaluation index, the limit state equation of tunnel allowable deformation in the soft rock was established. Finally, considering the spatial variability of elastic modulus of layered rock mass, a quantitative risk analysis of surrounding rock instability was realized by the Monte Carlo method and numerical simulation method. The engineering example shows that, this quantitative risk analysis method can effectively evaluate the safety state of the surrounding rock mass, and it can provide a theoretical guidance and technical support for the tunnel construction. This work could provide a reference for other layered rock mass tunnel construction.

Instability Control of Roadway Surrounding Rock in Close-Distance Coal Seam Groups under Repeated Mining

In order to solve the problems of roadway stability and easy instability under repeated mining of close-distance coal seam groups, the mechanism and control technology of surrounding rock instability under repeated mining were studied via indoor testing, field testing, physical similarity simulation experiment, and numerical simulation. The results show that the surrounding rock of roadway has low strength, low bearing capacity, and poor self-stabilization ability, and it is vulnerable to engineering disturbances and fragmentation. Affected by the disturbance under repeated mining, the roadway surrounding rock cracks are developed and the sensitivity is strong, and it is prone to large-scale loose and destroyed. The location of the roadway is unreasonable, and the maximum principal stress of the roadway is 3.1 times of the minimum principal stress, which is quite different. Thus, under a large horizontal stress, the surrounding rock undergoes long-range expansion deformation. On the basis of this research, the direction and emphasis of stability control of roadway surrounding rock under repeated mining of coal seam groups in close-distance are shown. A repair scheme (i.e., long bolt + high-strength anchor cable + U-shaped steel + grouting) is proposed, and reduces the risk of roadway instability.

Study on the Influence of In Situ Stress Distribution on the Stability of Roadway Surrounding Rock

In order to understand the instability characteristics of surrounding rock in the process of deep roadway excavation, a three-dimensional numerical model was established by FLAC3D to systematically analyze the influence of roadway surrounding rock stability under different in situ stress distribution forms, and the environmental coefficient of mining-induced stress η was defined, the larger the environmental coefficient of mining-induced stress is, the larger the surrounding rock stress environment is, and the range where the η coefficient is greater than 0.2 is called with the destruction-danger zone. When the initial vertical stress is maximum principal stress and minimum principal stress, by comparing the roadway along the middle ground stress direction and minimum or maximum in-situ stress direction, the variation characteristics of displacement, failure zone and failure hazard zone of roadway surrounding rock are obtained, which provides theoretical basis for the treatment of disaster accidents such as roadway surrounding rock instability and rock burst caused by deep high in-situ stress.

Roof Failure Mechanism and Control Technology of Large Section Open-Off Cut in Soft Rock Strata with Thin Thickness

The open-off cut is used for equipment installation of working face before underground mining, and its sectional size is larger than that of the mining roadway. Therefore, the stability of open-off cut surrounding rock determines whether the panel can be put into operation. To solve the roof instability of open-off cut in the Wanli No.1 coal mine, the roof failure mechanism of open-off cut under weak composite rock strata with thin thickness was studied by field monitoring, theoretical analysis, and numerical simulation. First, the characteristics of surrounding rock and the basic law of strata behaviors were obtained by detailed field monitoring. Afterward, FLAC3D numerical simulation and mechanical analysis were used to obtain the main mechanical control parameters of surrounding rock instability, and the existence of a soft interlayer above the roof is the main cause of roof instability. Based on this, the supporting parameters of the open-off cut were optimized and adjusted. The optimized parameters were applied to the adjacent 31207 open-off cut. The engineering practice showed that the optimized supporting parameters have an ideal control effect on roof stability.

Research on Failure Process and Precursor Information of Surrounding Rock of Deep Layered Roadway Based on Digital Image Correlation Method

In order to reveal the deformation and failure mechanism of surrounding rock of deep layered roadway, the deformation field of physical model experiment of surrounding rock of roadway with 45 degree dip angle is monitored by digital image correlation (DIC) method, and the deformation and failure process is analyzed. In view of the difficulties in monitoring and early warning of surrounding rock deformation and failure disaster, based on DIC, a new method for identifying the precursor information of surrounding rock failure called image coefficient of variation (ICV) is proposed, taking the displacement change and displacement change rate in the deformation field obtained from DIC analysis as calculation individuals, the ICV is obtained by calculating the whole field coefficient of variation (CV). The research shows that the DIC method is an effective non-contact deformation monitoring method in the tunnel roadway physical model experiment. Through comparative analysis, it is found that the variation trend of ICV curve can be used as the precursor information of roadway surrounding rock failure, which provides a new early warning index for roadway surrounding rock instability disaster.

Study on Monitoring Technology of Surrounding Rock in Deep Layered Roadway Based on Constant Resistance and Large Deformation Bolt

The stability monitoring of surrounding rock in layered roadway is an important method to ensure the safety of deep mining coal mine. In view of the actual situation of roadway stability monitoring, based on similarity theory, the deep roadway model is built and small-sized ordinary bolts and constant resistance and large deformation (CRLD) monitoring bolts are developed to detect the mechanical real-time information of surrounding rock in the model, the accurate measurement information of the displacement field and the force value of the bolts are obtained. According to the time series characteristics of the deformation field and the force curves of the monitoring bolts, the criterion of stability and instability pattern of surrounding rock in deep layered roadway are studied. The results show that the force monitoring method based on CRLD bolt is superior to displacement monitoring, the force monitoring curve of CRLD bolt can be divided into stable mode and instability mode, the characteristic of stable mode curve is rising slowly and the instability model with rapid rising and falling stages. The stability monitoring system of roadway with CRLD bolt is developed and applied in deep rock roadway, it has been proved by field practice that the early warning of surrounding rock instability of deep layered roadway is feasible.

Using the characteristics of infrared radiation b-value during the rock fracture process to offer a precursor for serious failure

The Infrared Radiation (IR) on the surface of a rock being loaded will change significantly during its failure process. To effectively monitor and predict rock failure by the IR method, in this paper, the Infrared Radiation Variation Temperature Field (IRVTF) of rock under uniaxial compression was identified. It was found that some IRVTF thermal images were characterized by obvious temperature rise in the later stage of rock loading. The temperature distribution characteristics of IRVTF were analyzed. It was observed that the temperature of the IRVTF exhibited the logarithmic linear distribution within the selected statistical temperature range, with an average coefficient of correlation greater than 0.96. On this basis, the index of “Infrared Radiation b-value (bIR-value)” of rock under uniaxial compression was defined for the first time, and the threshold for identifying its sudden change was proposed. It was found that the sudden changes of bIR-value could be observed with the generation of large-scale micro-cracks. The average amplitude of the bIR-value at first sudden change for all samples was 16.58 times of the mean amplitude before the sudden change on average, which indicated the significance of the bIR-value sudden change. The first sudden change of the bIR-value was 22.5 s ahead of the peak stress on average, which could be used as a precursor of rock failure and instability. The findings of the research can help with the monitoring and early warning of surrounding rock instability in mines, tunnels, and other geotechnical projects.

Acoustic Emission Characteristics and Damage Evolution of Rock under Different Loading Modes

To study the evolution of acoustic emission (AE) parameters and the differences in the fracturing and failure process of rocks under different loading modes, AE signals of marble were detected during uniaxial compression tests (UCTs), direct tensile tests (DTTs) and indirect tensile tests (ITTs) in this paper. Then, the temporal and spatial evolution of the AE parameters and damage development of rock under different loading modes were analyzed. The results showed that the sequence of total AE events and AE counts under different loading modes was UCT > DTT > ITT. In the DTT and ITT, the energy release of AE signals was concentrated at the peak stress and weakened rapidly afterward, whereas in the UCT, there were still a large number of AE signals accompanied by violent AE energy release during the postpeak stage. The generation mechanism of AE sources in rock and the corresponding failure modes were different under different loading modes. In the UCT, the multiple cleavage fractures were mainly caused by compression-induced fracturing. In the DTT, the single fracture surface was generated by tensile stress, whereas in the ITT, compressive-tensile stress was applied to the fracture surface. In addition, the stress levels at which the b-value and the spatial fractal dimension Ds of AE events decreased dramatically were consistent under the different loading modes, and the sequence was UCT < DTT < ITT. According to the changes in AE parameters during the whole process of rock deformation and failure, the first and second precursor points before failure were defined to distinguish the development of microfracture damage and failure processes in rocks under the different loading modes. The above results have certain significance for future studies on the monitoring of surrounding rock instability and failure prediction.

Distribution characteristics of the advance abutment pressure in a deep stope

Abstract The advance abutment pressure on the coal mining face of a deep stope plays a decisive role in the surrounding rock instability, rock burst and coal-gas outburst. Based on distributed optical fiber sensing technology and theoretical analysis of the stress state, the distribution characteristics of the advance abutment pressure were tested. Taking the 61103 coal mining face of a mine in Inner Mongolia as an example, downhole measurement research was carried out. Based on the monitoring data, a two-dimensional (2D) geological model for the distribution of the advance abutment pressure during the mining process was developed. To verify the validity of the model, a verification borehole was located on the 61201 coal mining face of the same mining area, the two borehole datasets were comprehensively analysed and the model was corrected. The modified model shows that the influence scope of the advance abutment pressure is 0–26.5 m, the position of the peak stress is 5.1–6.4 m in front of the coal mining face and the peak stress can reach 21.98–23.98 MPa. The stress concentration coefficient is 1.22–1.91 and the vertical influence depth is 13–14 m. Moreover, the advance abutment pressure exhibits obvious zoning in the transverse direction, and the stress at the same position in front of the coal mining face will undergo a process of steady-slow increase-significant increase-peak-decrease. The technical parameters obtained from this study can be used for other coal mining faces in the area. In addition, the testing result of the abutment pressure from the distributed optical fiber sensors is effective and can provide a new research method for mine safety monitoring.

Stability evaluation of flat large-span cavern in jointed rock mass

The excavation of the rock mass will have a great influence on the characteristics of the caverns under extremely bad geological conditions. According to the characteristics of the engineering geology, several groups of the caverns are designed to be excavated in the rock mass with three sets of through joints. Based on the summary of the failure mechanism and the failure pattern, the classification method of the failure modes is done for the flat large-span underground cavern with three sets of through joints. The classification method takes into account the control factors of the surrounding rock instability, failure mechanism, distribution of the disturbed zone, and characteristics of the rock weakness, etc. The failure modes of the flat large-span caverns contain six types: the relatively stable, the partial area falling, and the partial collapse of vault, the zigzag failure of rock stratum, the V-shaped failure, and the large-scale collapse. The change law of the vertical stress is first increasing and then decreasing for the surrounding rock in the disturbed zone. The horizontal stress shows an exponential change trend with an increase in the distance from the vault. Influenced by the failure degree of the caverns, the change law of the vertical displacement for the vault can be divided into two types. The horizontal displacement decreases with an increase in the distance from the sidewall, and there is an exponential relationship between them.