To effectively control the large deformation of the surrounding rock under complex conditions, it is often necessary to apply prestress to anchor cables. However, due to the influence of surrounding rock deformation, mining disturbance, and strong impact, anchor cables are often in a dynamic and static coupling stress state. Therefore, it is crucial to study the dynamic and static coupling mechanical characteristics of anchor cables. Based on this, the self-developed dynamic and static coupling test equipment is developed. The dynamic and static coupling mechanical test of anchor cables is conducted. Test results indicate that the energy absorption for anchor cables under the initial load of 350 kN decreased by 69.8% compared to the condition without initial load, and the energy absorption efficiency increased by 6.6 times. The increase of initial load can improve its energy absorption efficiency, but it can also lead to a decrease in its energy absorption. The energy absorption and energy absorption efficiency shows a bilinear variation law with the increase of initial load. On this basis, the energy absorption calculation formula and the support design model of the anchor cable are established. It provides new ideas for the safety control of dynamic disasters in deep engineering.

Deep underground engineering often faces complex conditions such as high stress and extremely soft rock. Due to the problems of heavy self-weight, insufficient bearing capacity and low construction efficiency, the traditional steel arch cannot meet the requirements of safe and efficient construction of underground engineering in deep and complex conditions. Based on this, a new lightweight high-strength arch support technology is developed, and full-scale mechanical tests of lightweight high-strength arch and U36 arch are carried out. The test results show that compared with U36 arch, the initial sliding resistance and ultimate bearing capacity of lightweight high-strength arch are increased by 37.6% and 20.1% respectively with 33.3% weight reduction. The new support can provide higher support strength for the surrounding rock of deep roadway while reducing the construction strength. On this basis, a full-scale numerical model of the lightweight high-strength arch is established. The influence mechanism of the lightweight high-strength arch bearing performance is clarified. The bearing capacity calculation equation of the lightweight high-strength arch is established. Taking the typical deep high stress mine as the engineering background, the field application of lightweight high-strength arch is carried out.

Accurate measurement of the compressive strength of rock and the degradation in its structural features is essential for the rational design of supports for rock masses in underground engineering. Traditional methods involving core sampling and laboratory testing struggle to meet the timeliness requirements of large-scale underground construction and intelligent development. In this study, an on-site drilling testing method for determining compressive strength and deterioration properties is established. An intelligent drilling explore system of geology (IDE system) is deployed for on-site drilling tests. The results show that the drilling parameters clearly respond to fracture zones and rock layer interfaces. When comparing the compressive strength values acquired using the IDE system with those of the laboratory tests, the average percent difference is 10.42%. The average rate of strength degradation for the structural features is 75.47%. The IDE system can effectively test the compressive strength of rock and its degradation on site. This study provides theoretical and technical support for on-site measurement and intelligent perception of surrounding rock parameters.

Dynamic hazards frequently occur in the gob-side roadway with coal pillars in deep coal mines. At present, there is a lack of systematic research on the impact failure mechanism of anchored sandstone roof with mining-induced stress and the effective control method in deep coal mines. Therefore, by conducting true-triaxial impact test experiment, this paper analyzed the impact failure mechanism of anchored sandstone under the different dynamic stress amplitudes and support material number. The experiment results showed that the larger the dynamic stress amplitude, the lower the peak strength and the earlier the impact failure time. Also, support material number can increase the peak strength of anchored rock specimens and greatly reduced the risk of impact failure. The cooperative control method of releasing roof stress and absorbing energy support in deep coal mines were proposed, and the effectiveness of this method was verified in a typical deep coal mine. Compared with traditional method, the results of using this novel control method showed that the maximum abutment stress was reduced by 20.66%, the total energy of micro-seismic events was reduced by 41.16%, and the maximum support force of absorbing-energy anchor cable was increased by 50.62%. This cooperative control method can provide guiding significance for deep safe mining.

Rockburst is a common disaster in deep underground engineering, which seriously impacts project construction safety. Understanding its causes and burst resistance mechanism is of significance for rockburst prevention and mitigation. We developed a new type of high strength, large elongation, and strong energy-absorbing material, and conducted comparative tests on both basic and anchored rock specimens with such material. We analyzed the rockburst process, energy release and peak stress of the rock, and force and deformation withstood by the energy-absorbing bolts. The experimental results show that the energy reduction rate of the rocks reinforced by energy-absorbing bolts is more than 80%, compared with that of the basic rock. The force exerted on the energy-absorbing bolts increases suddenly when the rockburst occurs, and the strength utilization rates of the energy-absorbing bolts under strain rockburst and impact rockburst conditions are 73.3% and 61.2%, respectively. Rockburst also causes non-uniform shear deformation of the anchor bolt. Based on the rockburst energy criterion, the peak stress of the anchored rock is 2.2 times and 2.5 times the uniaxial compressive strength of the rock, respectively, under strain rockburst and impact rockburst conditions. The energy required for rockburst is 396.0 and 478.4 kJ/m(3), respectively. The energy-anchoring bolts can effectively reduce the likelihood of rockburst. The results can provide a reference for support design for burst-prone rock in underground engineering.

Mining-induced seismicity is a distinct category of anthropogenic earthquakes, with their incidence increasing with mining depth. This study is grounded in mining-induced seismic data and technical solutions from typical mines in China. Mining-induced seismicity is defined as all detectable surface seismic motions within mining areas due to deep mining. This study first delves into the mechanisms of overburden breaking-type and fault slip-type mining-induced seismicity. The driving force is regional stress redistribution and concentration in the overlying strata. An integrated underground-surface monitoring network based on 4G/5G and GPS synchronisation is presented. The system’s data volume increased by 55.65%, and its validity increased by 198.29%. A hybrid algorithm integrating genetic algorithms with the Powell algorithm is developed to address challenges in source localisation. An evaluation model for monitoring efficiency is established, based on the Improved Non-Dominated Sorting Genetic Algorithm II. The evaluation of prevention technologies indicates that deep-hole blasting achieves significant efficacy by “replacing one major event with multiple minor events”. Overburden separation grouting has limited impact on energy release, and hydraulic fracturing remains inconsistent due to uncontrollable flow behaviour. This study contributes to an in-depth understanding of mining-induced seismicity, thus providing valuable references in the field of coal mine dynamic disasters.

The “anchor spray + arch” support system is widely used in tunnel design. However, under high in situ stress and large-span conditions, insufficient rock self-stabilization causes loads to be carried mainly by the arch and spray layer, leaving bolt capacity underutilized and compromising the synergistic interaction among support components, thereby increasing failure risk. To enhance rock self-stabilization, high prestress should be applied to bolts. Based on this, a cooperative control theory of high-prestressed supporting system was proposed. A high-strength constant-resistance bolt enabling high prestress application was developed, along with a numerical constitutive model of the active–passive support system. Numerical simulations and geomechanical model tests on the cooperative control of the high-prestressed support system for tunnels were conducted. The active regulation mechanism of prestress on surrounding rock load distribution was revealed. As prestress increased from 0 to 100 kN, the load proportion carried by the active support rose from 13.75 to 48.22

Accurate detection of the rock mass mechanical parameters and structural characteristics is essential for underground construction engineering and support design. As deep underground engineering activities encounter increasingly complex geological conditions, traditional methods for testing the rock mass parameters and structural characteristics are becoming inadequate for real-time, precise acquisition of on-site geological data. Digital drilling technology, based on field drilling equipment, provides an effective approach for real-time acquisition of in situ rock mass mechanical parameters and structural characteristics. In this study, a developed rotary cutting test (RCT) system for rock mass and an anchor bolt analytical drill bit are employed. Digital drilling tests are conducted on rock masses of varying types and strengths, and the relationships between the drilling parameters and rock compressive strength are analysed. Additionally, a rock compressive strength inversion model, DP-RCSs, is established based on the drilling parameters, and the average difference rate of the inversion model obtained from the verification test is 7.43%. Building on these developments, an intelligent detection anchoring (IDA) system for underground engineering is developed. Digital drilling model tests are performed on layered rock mass, and results show that the average difference rate for rock compressive strength measurements performed using the IDA system is 10.89%. The average difference rate in identifying structural plane positions is 1.69%. Furthermore, a field test of the IDA system is carried out in a deep coal mine roadway. The average difference rate of compressive strength detected while drilling is 12.21%, and the coal-rock interface is identified through drilling testing. This study provides a theoretical and technical foundation for the real-time in situ detection of surrounding rock parameters for deep underground engineering.

The accurate and real-time acquisition of rock strength parameters is very important for underground engineering construction. It is difficult for traditional laboratory testing methods to restore the rock strength properties in the field environment. The drilling test method provides a new way to obtain rock strength parameters in situ. However, the influence of confining pressure on drilling parameters and rock strength parameters should be considered under the conditions of high stress in deep surrounding rock. Based on this fact, digital drilling tests of specimens with different strengths under true triaxial conditions are carried out in this paper. The response rules of drilling parameters and cutting energy to changes in confining pressure and the response rules of cutting energy to variations in rock strength are clarified. An inverse model of equivalent compressive strength of rock while drilling is established, and the strength strengthening efficiency of 5–25 MPa confining pressure on different types of specimens is determined. Based on the above research, this study proposes a test method of surrounding rock strength while drilling under true triaxial conditions, providing a theoretical foundation for the in situ measurement of surrounding rock mechanical parameters in underground engineering.

ObjectiveAiming at the problems of unbalanced magnetic pull (UMP) and low structural strength of high-speed rotor in the operation of permanent magnet assisted magnetic gear, relevant research was conducted.MethodsFirstly, the phase tuning method was used to study the influence of different transmission ratios on the UMP. Then, according to the structural characteristics for the high-speed permanent magnet rotor of the magnetic gear, the analytical solution of the rotor strength was obtained by using the equivalent mass ring method. Finally, taking reducing the maximum stress and ensuring a certain output torque as the optimization objectives, the multi-objective optimization was performed on the relevant parameters of the magnetic isolation bridge and magnetic barrier.ResultsThe analysis results indicate that a transmission ratio where the maximum common divisor of the number of magnetic blocks and the number of poles of the low-speed rotor GCD=2 can effectively decrease UMP. The relative error between the analytical solution of the maximum stress obtained from the equivalent mass ring method and the result of finite element simulation is less than or equal to 1%, validating the accuracy of the analytical method. Through the optimized design, the maximum stress on the rotor is significantly reduced.

With the development of underground engineering to the deep, deep engineering construction often faces complex conditions such as high stress and extremely soft rock. Due to the problems of heavy self-weight, low construction efficiency, and insufficient bearing capacity, the traditional steel arch is difficult to meet the requirements of safe and efficient construction in deep underground engineering. Therefore, the new high-strength and high-efficiency arch (HSE arch) is developed, and the full-scale mechanical tests of the HSE arch and the U36 arch under different load distribution forms are carried out. The deformation failure characteristics and yielding bearing characteristics of the HSE arch and the U36 arch are clarified. The influence mechanism of different load distribution forms on the bearing capacity of the arch is revealed. The test results show that compared with U36 arches, the bearing capacity of HSE arches is increased by more than 13.6% under the condition of 33.3% weight loss. The HSE arch can still provide a higher support force while reducing the weight and increasing the construction efficiency. Compared with the uniform loading condition, the bearing capacity of the HSE arch and the U36 arch decreased by more than 40.0% and 42.0% respectively, under the bias loading condition. The bearing capacity of the arch under the bias loading condition is lower, and it is more prone to instability failure. On this basis, the mechanical model of the HSE arch with arbitrary segment numbers and unequal stiffness is established. The theoretical calculation method of the bearing capacity is proposed. The difference rate between the theoretical calculation results and the indoor test results is within 7%, which verifies the effectiveness of the theoretical calculation. The field application of the HSE arch is successfully carried out in the typical deep roadway.

The elastic modulus of rock mass is a fundamental parameter for the surrounding rock stability analysis and the support scheme design. The traditional testing methods are mainly conducted through indoor experiments, which require further research for in-situ testing of rock mass elastic modulus. This article conducts multi type rock mass digital drilling experiments based on the intelligent rotary cutting testing system for rock masses. The response law of drilling parameters to elastic modulus has been clarified. And a rock rotational ratio energy that integrates four types of drilling parameters is proposed. The rock elastic modulus prediction models (RD-Ei models) are established. The experimental results show that the average testing errors of the model based on drilling pressure, drilling torque, and rotational ratio energy are 21.04%, 18.84%, and 6.44%, respectively. On this basis, the intelligent drilling explore system of geology is used to carry out rock drilling experiments. The identification of rock interfaces and testing of elastic modulus can be achieved. This study lays a theoretical foundation for real-time quantitative measurement of the surrounding rock elastic modulus on site.

During the mining of lower coal seams beneath underground reservoirs, the interlayer rock mass undergoes deformation and failure due to mining-induced disturbances. Crack propagation and interconnection pose significant threat to reservoir safety. Taking the working faces of 2-2 and 5-2 coal seams in Shenhua Daliuta Mine as prototypes, this study conducted physical similarity simulation experiments to analyze stress and crack evolution patterns in interlayer rock masses during lower seam extraction. We classified fractures into trace lines and segment lines to investigate the evolution of their topological and geometric characteristics. A three-phase diagram of fractures was established based on average connection numbers to quantitatively evaluate connectivity. Results indicate that as mining advanced, the interlayer rock mass experienced a nonlinear increase in pressure, followed by sudden pressure relief and stress fluctuations. Cracks continued to expand until reaching a maximum aperture before closing. The number of nodes in trace lines and segment lines increased with mining advancement, with the longest trace line typically being a horizontal bedding separation fracture. During evolution, fracture networks predominantly propagated as connected clusters while maintaining spatial continuity. This study provides a method for calculating the safety distance of coal seam mining beneath underground reservoirs and evaluating the safety of interlayer rock masses in coal mines.

The geological conditions of deep engineering are complex and are impacted by high stress and excavation disruption, which can lead to engineering disasters like severe surrounding rock deformation and roof collapse. The real-time, accurate measurement of rock mechanical parameters is a necessary condition for analysing the stability of deep surrounding rock and preventing disasters. In this paper, drilling tests are conducted with an intelligent detection anchoring system for underground engineering. An investigation is conducted into the correlation and change law between drilling parameters and the elastic modulus E, and a testing model for E while drilling, is proposed. According to the test results, the average difference in E is 0.43 GPa, and the average error rate is 5.49%. On this basis, an in situ zoning approach for surrounding E while drilling is established, and field drilling experiments are carried out for deep surrounding rock. A three-dimensional (3D) zoning and mapping relationship model is established for the equivalent elastic modulus Eeq of the surrounding rock, enabling in situ testing of the Eeq of the surrounding rock and effective identification of rock interfaces and fracture zones. While constructing the boreholes for rock bolts and anchor cables, the drilling testing method established can constantly and quickly determine the E of rock in real-time. This work provides theoretical and technical support for the on-site, real-time acquisition of rock mechanical parameters required for preventing and controlling failures of surrounding rock in underground engineering.

In response to the difficulty of controlling the stability of the surrounding rock in deep mining chambers, taking a typical deep coal mine pump chamber group as the engineering background, the deformation development mode, loosening range expansion law, and anchor (cable) bearing state of the chamber group under the influence of construction disturbance were analyzed. A compensation control method for deep chamber group excavation with the core of "stress compensation, grouting reinforcement, and excavation disturbance reduction" was proposed. To further verify the rationality of this method, large-scale geomechanical model tests and numerical simulation tests were conducted. The distribution and evolution of stress and displacement fields in the surrounding rock during the construction process of chamber groups were analyzed. The results show that compared with traditional control methods, using the compensation control method for deep chamber group excavation increased the stress of the shallow surrounding rock by 68.2% and reduced the surface displacement of the surrounding rock by 35.5%. Based on the above research, the field tests were conducted. The monitoring results showed that the maximum deformation of the surrounding rock was 115 mm, and the utilization rate of cable strength was 63.8%, achieving stability control of the deep chamber group.

The propagation of rock fractures is essential for maintaining engineering safety, yet traditional theoretical methods are burdened by challenges such as complex sample collection and lengthy prediction processes. To address these challenges, this study develops a deep learning model based on an adaptive moment estimation optimized convolutional long short-term memory neural network (Adam-ConvLSTM) to predict the evolution of rock fractures. We generated a rock fracture dataset through numerical simulation and then incorporated it into a machine-learning imagework to produce a predictive model. Initially, PFC2D numerical simulations were conducted on rocks with various pore defects under uniaxial compression, resulting in five sets of fracture propagation images. These images were processed using sliding window techniques to construct a foundational dataset. Considering the spatiotemporal correlations among different rock fractures, one dataset was used to train the Adam-ConvLSTM model, yielding an initial model that accurately predicts fracture propagation for a single rock dataset. Utilizing transfer learning, this initial model was adapted and independently fine-tuned for four additional datasets with varying pore defect sizes, resulting in four distinct predictive models. These models were integrated to form a more comprehensive predictive system. In practical applications, the comprehensive model uses structure similarity index measure to align test samples with the most similar images from the model, selecting the predictive model with the most similar images for forecasting fracture evolution. Comparative validation indicates that this comprehensive model outperforms traditional methods and basic deep learning algorithms in both prediction efficiency and accuracy. This model not only enhances the efficiency and precision of rock fracture evolution forecasting but also offers a practical approach for monitoring rock mass fractures, substantially enhancing engineering safety.

Under the influence of the upper coal pillars and dynamic pressure of coal mining, the roadway of the lower coal seam is prone to large deformation failure. In this paper, a novel control method and key technologies of automatically formed roadway (AFR) by roof cutting and confined concrete column in extremely close-distance coal seam are proposed. Furthermore, a numerical model is established to analyze the structure characteristics of overlying roof strata. Based on numerical results, the roof structure model of “voussoir beam of upper layer + short cantilever beam of lower layer” of this method is proposed. What’s more, the calculation equation of the roof bending moment and evaluation indexes is established, and the influence of different factors on roof stability control of AFR is studied. Finally, a field test is conducted to verify the effectiveness of this novel method. Field results were as follows: 1) The maximum and average support stress of working face obviously decreased; 2) The confined concrete column can provide high-strength support in dynamic influence zone; 3) The maximum deformation of AFR safety requirement can be met. This study can provide effective guidance for the application of this method in extremely close-distance coal seam.

To control the surrounding rock of a deep chamber group, this paper establishes a mechanical model for the disturbance effect of deep chamber group excavation, analyzes the stress evolution law of the surrounding rock during the construction process of chamber groups, and proposes a compensation control method for deep chamber group excavation with "stress compensation, surrounding rock reinforcement, and excavation disturbance reduction" as the core. Taking a typical deep coal mine pump room chamber group as the engineering background, evaluation indicators such as the characterization deformation control rate and the control rate of the unloading zone range of the surrounding rock were established. Numerical experiments were conducted to study the influence mechanism of the pre-tightening force, surrounding rock reinforcement coefficient, and excavation sequence on the stability of the chamber groups. A high pre-tightening force was applied to reduce the unloading range of the surrounding rock, and the parameters of the surrounding rock were strengthened to improve its self-carrying capacity. The excavation compensation control mechanism of optimizing the excavation sequence to reduce stress disturbance was verified through the large-scale geomechanical model test. Furthermore, based on the research results, a compensation control design method for deep chamber group excavations was proposed and successfully applied in a deep mine chamber group. The maximum deformation of the surrounding rock was 115 mm, verifying the effectiveness of the compensation control design method for deep chamber group excavations.

The surrounding rock is prone to large-scale loosening and failure after the excavation of shallow large-span caverns because of the thin overlying strata and large cross-section span. The rational design of bolt support is very important to the safety control of surrounding rock as a common support means. The control mechanism and design method of bolt support for shallow-buried large-span caverns is carried out. The calculation method of bolt prestress and length based on arched failure and collapsed failure mode is established. The influence mechanism of different influencing factors on the bolt prestress and length is clarified. At the same time, the constant resistance energy-absorbing bolt with high strength and high toughness is developed, and the comparative test of mechanical properties is carried out. On this basis, the design method of high prestressed bolt support for shallow-buried large-span caverns is put forward, and the field test is carried out in Qingdao metro station in China. The monitoring results show that the maximum roof settlement is 6.8 mm after the new design method is adopted, and the effective control of the shallow-buried large-span caverns is realized. The research results can provide theoretical and technical support for the safety control of shallow-buried large-span caverns.

High in-situ stress, extremely soft rock, large sections, and other complex geological conditions are faced in the process of tunnel construction, leading to difficulties in controlling the surrounding rock. To address the issue of deformation control of surrounding rock under complex conditions, the prestressed support system is widely used in engineering. Understanding the surrounding rock control effect of prestressed support and the mechanical characteristics of support structures is crucial for the control design of tunnel surrounding rock. Model tests can directly and truly reflect the interaction between rock and support structures, which is an effective means to study the above engineering challenges. However, there is no corresponding theory, method and technology in the simulation of prestressed support. In this paper, a similarity theory and equivalent simulation method of prestressed support are proposed. A key simulation technology for high prestress application is developed. Taking the large-section tunnel in eastern China as the engineering example, the geomechanical model test of high prestressed anchorage support for tunnels is carried out. The results show that the deformation and stress release of tunnel surrounding rock are significantly controlled by high prestressed support. The average deformation control rate lambda D of the surrounding rock of the vault and hance is 51.29 % and 66.64%. The average stress release control rate lambda S of the surrounding rock of the vault and hance is 35.78 % and 18.47 %. According to the mechanical analysis of support structures, the high prestressed support can improve the mechanical characteristics between support structures. After applying high prestress, the axial force of bolts is increased by an average of 5.30 times, and the stress of the arch and spray layer is decreased by an average of 42.94 % and 42.73 %, respectively. Furthermore, we perform a laboratory test of different prestress application methods. The prestress conversion effect of different application methods and the prestress retention effect of bolt tension are investigated. Based on the above research, the field design and application of high prestressed support for surrounding rock are conducted.