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The accurate measurement of coal seam gas content is essential for several aspects of deep coal mining, including disaster management, resource allocation and sustainability. However, obtaining in-situ coal samples while preserving gas content under challenging conditions, such as high stress, temperature fluctuations, and drilling fluid environments, remains a significant challenge. To overcome this difficulty, we present an innovative insitu pressure- and gas-preserved coring tool specifically designed for deep coal mining applications. This device enables the collection of coal seam samples under in-situ conditions while ensuring that both pressure and gas content are preserved, thereby preventing gas escape during sample transfer and providing more accurate parameters for evaluating coal and natural gas reserves. In the demanding environment of deep coal seams, the performance of the pressure-preserved chamber of the corer relies on the reliability of its remote triggering mechanism. The presence of drilling fluid introduces medium resistance, which can impair the triggering process--an issue largely overlooked in previous research. Herein, we propose a robust method to calculate remote triggering forces within liquid media and optimize its key parameters to improve operational stability. Laboratory tests and field validations in coal mining environments are conducted, which confirm the effectiveness of the optimized design and demonstrate the tool's practical applicability. This study offers valuable insights into addressing key challenges in deep coal reservoir exploration and gas resource preservation. [ABSTRACT FROM AUTHOR]

Deep coal mining, in addition to facing the three “high” and one “disturbance”, faces a high-stress corrosion environment formed by underground corrosive atmosphere, high stress, strong disturbance, and other factors. Taking the deep mines in the Juye Coalfield, Shandong, as the engineering background, the high-stress corrosion characteristics and influencing factors of deep mining are analyzed. Experimental and theoretical research is conducted on the anchor rod material to reveal the degradation mechanism of anchor rod strength under high-stress corrosion environment, and a theoretical model for the degradation of rod strength is derived. The results are as follows: ① through the SEM, corrosion SSRT, and accelerated corrosion experiments on rod materials, it is found that the corrosion causes a decrease in the elongation at break and fracture time of rod materials by 8.22% and 8.34%, respectively, compared to inert conditions. With the increase of corrosion rate, the material failure mode transitions from ductile to brittle fracture, and the localized corrosion type gradually evolves from pitting to uniform corrosion. ② Combining experimental results with electrochemical theory and non-uniform corrosion models, introducing the Pruckenr temperature influence factor, a time-varying model for the degradation of rod material strength is derived. The in-depth investigation of the impact of service environment temperature T, rod cross-sectional radius R, and material room temperature corrosion rate Vh,298K on the service life of the rod reveals that R is linearly positively correlated with service life. The relationship between T and Vh,298K with service life shows that the degree of service life decay decreases as the levels of influencing factors increase. ③ Combining experimental results with the energy absorption principle of anchor rods, the limit energy absorption capacity of roadway anchor rod support is calculated and analyzed. The results show that as the corrosion rate of the anchor rod increases, the energy absorption of the anchor rod decreases, the maximum seismic resistance decreases, and the risk of roadway dynamic instability disasters increases.

The grouted bolt, combining rock bolting with grouting techniques, provides an effective solution for controlling the surrounding rock in deep soft rock and fractured roadways. It has been extensively applied in numerous deep mining areas characterized by soft rock roadways, where it has demonstrated remarkable control results. This article systematically explores the evolution of grouted bolting, covering its theoretical foundations, design methods, materials, construction processes, monitoring measures, and methods for assessing its effectiveness. The overview encompassed several key elements, delving into anchoring theory and grouting reinforcement theory. The new principle of high pretensioned high-pressure splitting grouted bolting collaborative active control is introduced. A fresh method for dynamic information design is also highlighted. The discussion touches on both conventional grouting rock bolts and cable bolts, as well as innovative grouted rock bolts and cables characterized by their high pretension, strength, and sealing hole pressure. An examination of the merits and demerits of standard inorganic and organic grouting materials versus the new inorganic–organic composite materials, including their specific application conditions, was conducted. Additionally, the article presents various methods and instruments to assess the support effect of grouting rock bolts, cable bolts, and grouting reinforcement. Furthermore, it provides a foundation for understanding the factors influencing decisions on grouted bolting timing, the sequence of grouting, the pressure applied, the volume of grout used, and the strategic arrangement of grouted rock bolts and cable bolts. The application of the high pretensioned high-pressure splitting grouted bolting collaborative control technology in a typical kilometer-deep soft rock mine in China—the soft coal seam and soft rock roadway in the Kouzidong coal mine, Huainan coal mining area, was introduced. Finally, the existing problems in grouted bolting control technology for deep soft rock roadways are analyzed, and the future development trend of grouted bolting control technology is anticipated. [ABSTRACT FROM AUTHOR]

China's coal mines are mainly underground mines, and a large number of roadways have to be excavated underground. It is of great significance for coal mine production to adopt safe and reasonable roadway support methods. In the process of roadway excavation, the rock stratum is inclined and the roadway pass through the layer. Since the surrounding rock conditions of the roadway passing through the layer are more complicated, it is easy to cause deformation of surrounding rock, failure and floor heave, which makes the support work difficult. In order to solve this problem, the mechanical properties of roadway surrounding rock were tested and the failure of roadway surrounding rock was analyzed using the +260 horizontal centralized transportation roadway in Changcheng No.2 mine. The surrounding rock of the roadway was divided into 8 regions, and the stress analysis of the surrounding rock in different regions was carried out. It is found that the left shoulder pit, the right side and the floor of the roadway are prone to damage. The influence of the lateral pressure coefficient, the rock dip angle and the lithology on the failure of the roadway surrounding rock was analyzed by Mohr-Coulomb failure criterion, and the specific failure range of the roadway surrounding rock was obtained. The support optimization design of the roadway was carried out, and the weak area of the surrounding rock was reinforced. The deformation monitoring of roadway surrounding rock after support optimization was carried out. The field monitoring results show that after the optimized support, the displacement of the roof and floor of the roadway section and the two sides are reduced by 43.6% and 40.8% respectively compared with the original scheme, and the deformation of the surrounding rock also shows a trend of gradual stability, and the surrounding rock of the roadway is effectively controlled. The research can provide a new way for the stress and failure analysis of the surrounding rock of the inclined rock roadway. [ABSTRACT FROM AUTHOR]

China’s coal mines are mainly underground mines, and a large number of roadways have to be excavated underground. It is of great significance for coal mine production to adopt safe and reasonable roadway support methods. In the process of roadway excavation, the rock stratum is inclined and the roadway pass through the layer. Since the surrounding rock conditions of the roadway passing through the layer are more complicated, it is easy to cause deformation of surrounding rock, failure and floor heave, which makes the support work difficult. In order to solve this problem, the mechanical properties of roadway surrounding rock were tested and the failure of roadway surrounding rock was analyzed using the +260 horizontal centralized transportation roadway in Changcheng No.2 mine. The surrounding rock of the roadway was divided into 8 regions, and the stress analysis of the surrounding rock in different regions was carried out. It is found that the left shoulder pit, the right side and the floor of the roadway are prone to damage. The influence of the lateral pressure coefficient, the rock dip angle and the lithology on the failure of the roadway surrounding rock was analyzed by Mohr-Coulomb failure criterion, and the specific failure range of the roadway surrounding rock was obtained. The support optimization design of the roadway was carried out, and the weak area of the surrounding rock was reinforced. The deformation monitoring of roadway surrounding rock after support optimization was carried out. The field monitoring results show that after the optimized support, the displacement of the roof and floor of the roadway section and the two sides are reduced by 43.6% and 40.8% respectively compared with the original scheme, and the deformation of the surrounding rock also shows a trend of gradual stability, and the surrounding rock of the roadway is effectively controlled. The research can provide a new way for the stress and failure analysis of the surrounding rock of the inclined rock roadway.

Long-term coal mining has created unique microbial communities in deep coal mines. Revealing the microbial community structure and metabolic functions in the underground environment can contribute to a better understanding of the coal mine ecosystem. In this study, we collected underground sediment samples from producing mines in eastern China at mining depths of −400 to −1100 m and performed high-throughput sequencing. Results showed that most of the genera in the underground sediment can degrade organic matter, such as polycyclic aromatic hydrocarbons (PAHs), benzene, toluene, and xylene, etc. The dominant genera in the underground sediment were Hydrogenophaga, Thauera, Pseudomonas, Rhodobacter, and Dietzia. Samples were divided into coal roadway (CR) and rock roadway (RR) groups according to the sampling location. The microbial community structure differed significantly (p < 0.05) between these two groups of samples, with the distribution of main genera in the CR group samples showing a negative correlation with Cu and a positive correlation with temperature. The samples from the CR and RR groups were significantly different (p < 0.05) in their metabolic functions, including membrane transport, metabolism of other amino acids, folding, sorting, and degradation. Microorganisms in the RR group samples showed high resistance to heavy metals, while microorganisms in the CR group had higher degradation functions of organic pollutants. Bugbase phenotypic predictions indicated a high potential pathogenicity of microorganisms in coal mine sediment, which was mainly contributed by the genera Hydrogenophaga, Pseudomonas, Geothermobacter, and Methylophaga, etc. This study deepens the understanding of microbial communities in deep coal mine environments; however, the organic contamination and biological health risks of underground environments require extensive attention. [ABSTRACT FROM AUTHOR]

Copyright of Journal of Central South University is the property of Springer Nature and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

Workers in deep coal mines are often exposed to hyperthermal conditions, which is a health and safety hazard. Although many novel approaches have been proposed in recent years, air distribution optimization combined with thermal analysis based on existing cooling systems has rarely been conducted. Here, heat dissipation in the deep mine environment was estimated based on field measurements and numerical simulations. A pilot study was conducted to validate the performance of the multistage cooling system and optimize the distribution ventilation. The results show that the “2 + 3” layout (Two air cooler outlets are in the machine lane, and three air cooler outlets are in the open-off cut) was superior to other schemes in terms of the temperature field and energy consumption. In addition, the outlet air temperature should be set to ≤20 °C, and the air volumetric flowrate of the longest outlet should be > 400 m3/min to improve the thermal environment of the wind lane. This effectively reduces the average temperature of the entire working face by 4.6–5.1 °C compared to the situation without cooling measures, thereby increasing the comfortable zone of the working face (

Owing to the high dimension and nonlinear characteristics of gas outbursts in deep coal mines, an intelligent evaluation method for systematically screening and integrating gas data in deep coal mines is proposed herein to effectively identify coal and gas outbursts in deep mines. A rough set improved using a genetic algorithm is introduced to reduce the dimension of complex data pertaining to deep coal mine gas to determine the main control index of deep coal and gas outbursts. Subsequently, the initial weight and threshold of a back propagation (BP) neural network are optimized by combining the characteristics of parallelism and robustness of the genetic algorithm. An adaptive optimization of BP neural network by genetic algorithm back propagation (GA-BP) model is established to identify gas outburst in deep coal mine reasonably. Compared with the standard BP neural network, data fitting shows that the method can significantly improve the detection accuracy of deep coal and gas outburst while improving the speed of disaster identification, as well as improve the efficiency of disaster identification, thereby increasing the risk identification accuracy of deep coal and gas outburst to 90%. This not only provides a new method for the scientific evaluation of deep coal and gas outburst risk, but also an important reference for the scientific evaluation of other high-dimensional and nonlinear fields. [ABSTRACT FROM AUTHOR]

The pipe jacking method has high construction safety, fast tunneling speed, convenient and flexible construction. However, if it is used in deep strata of coal mines, it will face challenges such as high ground stress and complex geological conditions. Taking the pipe jacking tunneling roadway in Zhangji Coal Mine as the engineering background, the similar model test of pipe jacking is carried out to reveal the bearing mechanism of pipe, and the design parameters of pipe are determined to the test results. Combined with the monitoring data of the construction site, the construction experience is summarized as follows: pipe deflection and slurry leakage are the main factors causing pipe damage. Under the premise of rectification and sealing work, the pipe jacking method has great advantages in coal mine gas and water disaster control roadway excavation engineering.

Due to the gradual depletion of shallow mineral resources at present, mines are now gradually entering the deep mining stage. To promote the safe and efficient green mining of deep coal resources and sustainable energy development, and to improve the production efficiency of paste filling mining, the research group has performed this study on the green filling mining technology and application of the working face. Taking working face 1241 (3) of the Xieqiao coal mine as the engineering background, the selection and experiment of filling materials were carried out, and the gangue, fly ash and cement produced by the Xieqiao coal mine were used as the filling aggregate. Next, the strength changes before and after paste filling was obtained by theoretical calculation. The strength at the early stage of filling was no less than 0.13 MPa, and that at the late stage of filling was no less than 2 MPa. Based on previous experimental research and theoretical calculation, the mixing pumping process of paste material ratio and the gangue crushing process were determined, and the filling pipeline system was designed. Then, based on the traditional coal mining technology, a filling mining technology of working face was designed and optimized. The field application of the research results shows that after the goaf of the working face had been filled, the ground pressure behavior of the coal wall of the working face was significantly weakened, and the stability of the surrounding rock of the working face was effectively controlled. Therefore, the method achieved good results, effectively controlled the stability of surrounding rock in goaf, and provided a theoretical basis and data support for realizing safe, efficient and green mining of deep coal resources. The results of this study bear important significance and application value.

With the development of coal resource mining at depth, rock burst and other disasters faced by deep mining are more and more common. However, deep mining still needs to face the complex geological environment of high stress. This study first analyses the structural characteristics and mechanical model of gob-side entry retaining and puts forward the corresponding control technology. Based on that, the technology of gob-side entry retaining by directional roof cutting is introduced, including pressure relief technology by directional roof cutting and constant resistant and large deformation anchor cable support technology. At the same time, reasonable charging structure parameters are determined by borehole observation equipment, and support parameters are designed considering the roof lithology and impact-resistance ability. Through the field engineering practice, surrounding rock of gob-side entry retaining is effectively controlled by analysing the monitoring equipment, which lays a theoretical and practical foundation for the large-scale application of the technology of gob-side entry retaining by directional roof cutting in deep coal mines. [ABSTRACT FROM AUTHOR]

Abstract The stability control of surrounding rock for a large or super‐large section chamber group is a difficult technical problem in deep mining conditions, and this stability control has become one of the most important factors restricting the safety of a large‐scale coal mine. Based on the in‐site geological conditions of a super‐large section chamber group that functions for a coal gangues separation system in the Longgu Coal Mine, we first researched the stress, deformation, and failure characteristics of the in‐site chamber group surrounding rocks by using the FLAC3D software. Simulation results showed that the maximum vertical stress, deformation, and failure range of the surrounding rock for a super‐large section chamber group are larger than those of an ordinary section chamber. In addition, the roof subsidence and the plastic zone radius at the intersection are obviously larger than that of other parts, which increase by approximately 50.8% and 44.4%, respectively. Therefore, the chamber group intersection should be taken as the key area for surrounding rock control. Then, the influences of two key parameters of the chamber group, spacing and angle, were researched in detail to help in the design of the chamber group. Results show that when the chamber spacing is 80 m, the interaction begins to occur. When the angle is 70°, the stress of the surrounding rock and roof subsidence reach the minimum. Thus, the optimum chamber group parameters are determined to be a spacing of 80 m and an angle of 70°. Finally, from the perspective of chamber group stability, the optimal chamber group parameters can better meet the normal use for the super‐large section chamber group. This research provides a reference for the design of a super‐large section chamber group under the same or similar conditions and provides a strategy for controlling the surrounding rock.

Abstract In underground mines, high air temperatures in the summer months lead to an increase in inlet airflow temperatures. This leads to seasonal thermal pollution in the mines. This paper examines the dynamics and effects of seasonal variation in surface air temperatures and surrounding rock temperatures in deep coal mines. It also examines temperature variations in the main ventilation circuit, working face, and surrounding rock. The study results revealed that airflow temperatures were significantly affected by seasonal air temperature variations. The greater the distance was between the inlet and the wellhead of the ventilation shaft, the less the effect was on temperature. Moreover, slight temperature variations (1.0–3.0 °C) were observed between various points on the return route during the summer months. Airflow temperatures along the airflow inlet to the return route of the working face first decreased, but then increased. The temperature field of the surrounding rock increased gradually with increased distance between the mine roadway and inlet, with recorded rock temperatures as high as 40.53 °C. The radius of the heat-adjusting layer was between 28 and 33 m.

The characteristics and evolution of the geo-stress field are the research bases of overburden migration, roadway surrounding rock stability, and dynamic disaster prevention in deep coal mines. However, the present understanding of mine geological stress mainly depends on measurements obtained at a few discrete points, leading to obviously insufficient knowledge of the geo-stress field in the whole mining area. In this study, a novel optimal inversion analysis model of the geo-stress field was established based on multiobjective constraint technology, solid mechanics, optimization analysis, and stress inversion theory. Then, a 3D numerical simulation analysis model with large-scale general finite element software as the solver was constructed for the optimization of the boundary conditions of the geo-stress field. This simulation model identified the key links between the stress field boundary loads and the actual measurement data by directly solving the multiobjective optimization problem. Six on-site measurement points were selected at the Xinji No. 1 mine, and the magnitude and orientation of the geo-stress at these points were measured with the hollow inclusion strain gauge method. The distribution of the geo-stress and the main influential factors were then analyzed numerically. The results of the study are of great significance for investigating the geo-stress distribution. [ABSTRACT FROM AUTHOR]

Superhigh‐water backfill mining can mitigate rockburst risks in rockburst‐prone deep coal mines and reduce pollution caused by traditional cemented paste backfill. In this mining method, the width of the rib pillar between two longwall panels is important to maintain roadway stability, prevent rockbursts and water seepage. To design the width of rib pillars, this study firstly established a T‐shaped overlying strata structure model and analyzed source of stress that caused deformation of the T‐shaped model. Based on this model, the abutment stress in the rib pillar was determined. Then, the criterion for overall burst instability of the rib pillar was proposed according to the derived abutment stress. The limit equilibrium theory was applied to obtain the pillar plasticity which can be used as the criterion for water‐seepage prevention. The proposed approach was used to design the width of a rib pillar in Yineng Coalmine located in Shandong Province, China. The analysis of the microseismic monitoring results and borehole drill cuttings show that the designed rib pillar with a width of 10 m was stable without water seepage during mining, indicating the width design method proposed in this study is effective. [ABSTRACT FROM AUTHOR]

The stability control of surrounding rock for a large or super‐large section chamber group is a difficult technical problem in deep mining conditions, and this stability control has become one of the most important factors restricting the safety of a large‐scale coal mine. Based on the in‐site geological conditions of a super‐large section chamber group that functions for a coal gangues separation system in the Longgu Coal Mine, we first researched the stress, deformation, and failure characteristics of the in‐site chamber group surrounding rocks by using the FLAC3D software. Simulation results showed that the maximum vertical stress, deformation, and failure range of the surrounding rock for a super‐large section chamber group are larger than those of an ordinary section chamber. In addition, the roof subsidence and the plastic zone radius at the intersection are obviously larger than that of other parts, which increase by approximately 50.8% and 44.4%, respectively. Therefore, the chamber group intersection should be taken as the key area for surrounding rock control. Then, the influences of two key parameters of the chamber group, spacing and angle, were researched in detail to help in the design of the chamber group. Results show that when the chamber spacing is 80 m, the interaction begins to occur. When the angle is 70°, the stress of the surrounding rock and roof subsidence reach the minimum. Thus, the optimum chamber group parameters are determined to be a spacing of 80 m and an angle of 70°. Finally, from the perspective of chamber group stability, the optimal chamber group parameters can better meet the normal use for the super‐large section chamber group. This research provides a reference for the design of a super‐large section chamber group under the same or similar conditions and provides a strategy for controlling the surrounding rock. [ABSTRACT FROM AUTHOR]

In underground mines, high air temperatures in the summer months lead to an increase in inlet airflow temperatures. This leads to seasonal thermal pollution in the mines. This paper examines the dynamics and effects of seasonal variation in surface air temperatures and surrounding rock temperatures in deep coal mines. It also examines temperature variations in the main ventilation circuit, working face, and surrounding rock. The study results revealed that airflow temperatures were significantly affected by seasonal air temperature variations. The greater the distance was between the inlet and the wellhead of the ventilation shaft, the less the effect was on temperature. Moreover, slight temperature variations (1.0–3.0 °C) were observed between various points on the return route during the summer months. Airflow temperatures along the airflow inlet to the return route of the working face first decreased, but then increased. The temperature field of the surrounding rock increased gradually with increased distance between the mine roadway and inlet, with recorded rock temperatures as high as 40.53 °C. The radius of the heat-adjusting layer was between 28 and 33 m. [ABSTRACT FROM AUTHOR]

Affected by high ground stress, the surrounding rock control of a roadway is one of the most important factors restricting the utilization of deep resources. Therefore, it is necessary to propose a method to improve the stress environment of the deep-buried roadway and reduce its deformation. The article focuses on the 121,302 machine roadway in Kouzidong coal mine to analyze the large deformations of roadways near goafs (RNGs) in deep coal mines and reveal the mechanisms related to pressure relief via roof cutting. Through physical simulation, overburdened structures and the migration laws of RNGs in deep coal mines are studied, and the overburdened RNGs will eventually have a double short-arm “F”-type suspended roof structure. The superposition movement of the structure is the prime cause for the large deformation of the RNGs considered here. Artificial roof cutting can weaken the superposition effect of the double “F” structure and induce the roof to produce a new fracture. Meanwhile, sliding deformation along the fault line releases greater stress, and the cut roof can better fill the goaf. The stress distribution ratio between goafs and the coal pillar is improved. Here, a mechanical model of key block B’ (KBB’) is considered and the stability criterion of KBB’ is obtained. According to the theoretical calculation here, the stress of a coal pillar could be reduced by 19.14% when KBB’ is cut along the edge of the coal pillar in the 121,302 machine roadway. After engineering verification, the field observation result shows that the deformation of the 121,302 machine roadway is reduced by more than 50% after roof cutting.

The friction wear and thermal fatigue cracking of the brake shoe and friction-induced self-excited vibration (frictional flutter) of the disc brake can easily occur during emergency braking of a deep coal mine hoist with at high speed and with a heavy load. Therefore, tribo-brake characteristics between the brake disc and brake shoe during emergency braking of a deep coal mine hoist are investigated in the present study. Scaled parameters of the disc brake of a deep coal mine hoist are determined by employing the similarity principle. Friction tests between friction disc and brake shoe are carried out to obtain the coefficient of friction in the case of high speed and large specific pressure between the friction disc and brake shoe. Coupled thermo-mechanical finite element analyses of the brake disc and brake shoe are established to investigate temperature and stress fields of the brake disc and brake shoe during emergency braking, which is validated by the engineering failure case. Effects of braking parameters on flutter characteristics between the brake disc and brake shoe are explored by employing a double-degrees-of-freedom vibration mechanism model. The results show that the maximum temperature, equivalent Von Mises stress and contact pressure are all located at the average friction radii of contact surfaces of the brake disc and brake shoe during emergency braking. The cage crashing accident in the case of high speed and heavy load in a typical coal mine shows crack marks and discontinuous burn marks at central locations of brake shoe and brake disc surfaces, respectively, which indicates frictional flutter characteristics between brake disc and brake shoe. During emergency braking, flutter time duration decreases with increasing initial braking speed and damping parameter; the flutter amplitude and frequency of the disc brake increases with increasing normal braking load and stiffness, respectively.

Copyright of Coal Science & Technology (0253-2336) is the property of Coal Science & Technology and its content may not be copied or emailed to multiple sites or posted to a listserv without the copyright holder's express written permission. However, users may print, download, or email articles for individual use. This abstract may be abridged. No warranty is given about the accuracy of the copy. Users should refer to the original published version of the material for the full abstract. (Copyright applies to all Abstracts.)

This paper presents an integrated investigation of the ground stability of a deep gateroad with a 1 km burial depth based on a field test, case studies, and numerical modelling. In situ stress measurements and mechanical properties tests were first conducted in the test site. Then, the deformation behavior, stress and yield zone distributions, as well as the bolts load of the gateroad, were simulated using FLAC3D software. The model results demonstrated that the soft rock properties and high in situ stress were the main factors for the deep gateroad instability, and the shear slip failure induced by the high stress was the primary failure model for the deep rock mass. In addition, the unsuitable support patterns, especially the relatively short bolts/cables with low pre-tensions, the lack of high-strengthen secondary supports and the unsupported floor strata, also contributed to the gateroad instability. Subsequently, a new combined supporting strategy, incorporating longer bolts/cables, yielding ring supports, and grouting measures, was proposed for the deep gateroad, and its validity was verified via field monitoring. All these could be a reference for understanding the failure mechanism of the gateroad with 1 km burial depth. [ABSTRACT FROM AUTHOR]

Traditional intensive drilling (or grooving) technology in a roadway destroys the rock mass and supporting structure in the shallow anchoring area while implementing pressure relief. In Dongpang coal mine, it is difficult to avoid the failure of the supporting system and continuous large deformations of the deep roadways influenced by disturbance, despite the application of strong bolt (cable) supporting systems and grouting reinforcements. Therefore, this study proposes a novel "anchoring and pressure relief" synergistic supporting technique to eliminate the dilemma among the anchoring supporting and pressure relief. With strong anchoring supporting in the shallow surrounding rock, this novel technology realizes pressure relief by excavating large-diameter holes in the deep peak stress area, to enable active regulation of the move of the peak stress area to the deep surrounding rock without impairing the shallow anchoring area. The path of coal body rheology and stress move is cut off by large-diameter holes using this novel technology, achieving the purpose of absorbing energy and deformation, decreasing the stress applied on the supporting structure in the anchoring area. The novel technology effectively solves the control problems of continuous large deformation in deep roadway and has a good engineering application prospect. This study findings will add scientific value for the improvement of surrounding rock control of deep roadway. [ABSTRACT FROM AUTHOR]

• A numerical model is established to quantitatively characterize the sustainability of the geothermal recovery system in backfilled stopes of coal mines. • The overall geothermal energy capacity provided by the system is evaluated. • The sustainable heat extraction rate of the system is evaluated. • The benefit of groundwater flow to system sustainability is quantified. Deep underground mines offer a great potential as low-enthalpy geothermal resource to be adopted in direct regional heating. In this paper, a parallel horizontal ground heat exchangers (pHGHE) system is introduced to utilize the geothermal resource in deep coal mines, however, the sustainability of the system in a long-run performance has not been well addressed in the previous study. This work presented a quantitative evaluation on the long-term sustainability of pHGHE system installed in backfilled stopes of deep coal mines. A novel numerical model is applied to characterize the heat transfer occurring between the horizontal pipes and backfill stopes. The proposed model was first verified to an analytical solution to determine the model reliability and accuracy. The validated numerical model was further applied to a typical underground backfilled stope of Anju coal mine in China. The results demonstrated that the maximum sustainable specific heat extraction rate must not exceed 40 W/m to maintain a 50-year exploitation period. A total thermal capacity of 17 GWh could be realized during each extraction cycle under the current conditions, which could provide heating to more than 170,000 m2 of residential building area. Accordingly, sensitivity analysis was performed to identify the key parameters influencing system sustainability. In addition, the effects of groundwater flow on the thermal interaction and long-term sustainability of the pHGHE systems were examined. Groundwater flow can be beneficial to the long-term sustainability of the system. This work confirms the feasibility and sustainability of geothermal resource utilization in backfill stopes of underground coal mines. The insights gained in this study can provide technical guidance for the design of geothermal systems in similar coal mines. [ABSTRACT FROM AUTHOR]

In terms of the large deformation unstable failure and its control problem of deep dynamic pressure in gateroad, the mechanical model of circular roadway under dynamic pressure environment in deep mine was established. Also, the implicit boundary equation of plastic zone around deep dynamic pressure gateroad was deduced. Based on this, the morphology evolution rules of plastic zone in deep dynamic pressure roadway was deeply analyzed. Then the mechanical conditions for the formation of the class I and class II butterfly-shaped plastic zone were illustrated, and the definitions of malignant expansion of the plastic zone and its criticality were made. Finally, the failure mechanism of deep dynamic pressure gateroad was revealed. The results indicate that it is easier for surrounding rock to produce the butterfly-shaped plastic zone under dynamic pressure effect in mining than the normal conditions,and the development size of butterfly leaf, plastic damaged zone are greater. With the increase of dynamic pressure influence, the regional stress field of gateroad becomes extraordinary high stress field and the plastic failure of roof, floor and ribs expand to deep rock mass, which leads to the more serious deformation and formation of malignant expansion of plastic zone. Eventually that cause a large deformation failure of surrounding rock. For the gateroad design and supporting under deep dynamic pressure, how to avoid or reduce the impact of dynamic pressure, improve surrounding rock stress environment and decrease the failure depth of butterfly leaf must be all considered adequately, which are beneficial to the maintenance of gateroad. [ABSTRACT FROM AUTHOR]

An imperative task for successful underground mining is to ensure the stability of underground structures. This is more so for deep excavations which may be under significantly high stresses. In this manuscript, we present stability studies on two tunnels, a horseshoe-shaped and an inverted arch-shaped tunnel, in a deep coal mine in China, performed using the 3DEC distinct element code. The rock mass mechanical property values for the tunnel shapes have been estimated through a back-analysis procedure using available field deformation data. The back-analysis has been carried out through a pseudo-time dependent support installation routine which incorporates the effect of time through a stress-relaxation mechanism. The back-analysis indicates that the rock mass cohesion, tensile strength, uniaxial compressive strength, and elastic modulus values are about 35-45 % of the corresponding intact rock property values. Additionally, the importance of incorporating stress relaxation before support installation has been illustrated through the increased support factor of safety and reduced grout failures. The calibrated models have been analyzed for different supported and unsupported cases to estimate the significance and adequacy of the current supports being used in the mine and to suggest a possible optimization. The effects of supports have been demonstrated using deformations and yield zones around the tunnels, and average factors of safety and grout failures of the supports. The use of longer supports and floor bolting has provided greater stability for the rock masses around the tunnels. Finally, a comparison between the two differently shaped tunnels establishes that the inverted arch tunnel may be more efficient in reducing roof sag and floor heave for the existing geo-mining conditions. [ABSTRACT FROM AUTHOR]

Comprehensive gas control of high methane working face in more than 900 m depth has always been a worldwide problem. An similar material simulation experiment was carried out based on the geological conditions of 1112(1) face in a kilometer deep coal mine, where the maximum gas emission quantity was 124.87 m3/min. The results show that a gas enrichment area,which is fan-shaped ring zone with a width of 30 m, exists and is located at 10 times mining height from the roof of mined coal seam in the vertical direction and starts from 20 m between the gob-side roadway in the horizontal direction, and the stress variation range along the coal seam by mining is not more than 30 m out of the track roadway at 10 times mining height. Therefore, the multi-purposes roadway was constructed at about 20 m higher than the roof strata and about 35 m outside the track roadway. And then the new comprehensive gas control mode including three lanes with a face, one roadway with multi-purposes, joint control and continuous mining was put forward. Engineering practice indicated that the maximum of methane extraction amount and the concentration of single penetrating borehole was 9.84 m3/min and 98.5% respectively, the maximal gas drainage efficiency was 92.4%, the average volume fraction of gas in return airflow was 0.4%, the maximal output of coal was 10 324.6 t/d and the total gas extraction quantity is 24.572 million m3. The mode proposed can achieve a safe and efficient simultaneous mining of coal and gas. [ABSTRACT FROM AUTHOR]

With the continuous exploitation of deep coal resources, the long-term stability control of surrounding rock of deep high stress soft rock roadway has become an urgent support problem. Because the traditional surrounding rock deformation zoning theory does not consider the creep effect of rock, it is not suitable for the support design of deep soft rock roadway, and can not ensure the long-term safe production of coal mines. This work proposes a novel four-stage strain-softening model based on rock creep characteristics, reveals the deformation zoning mechanism of surrounding rock in deep high stress soft rock roadway, and proposes the determination method of actual deep roadway surrounding rock zoning type and boundary stress. Moreover, taking Pan Yidong Coal Mine in China as an example, the feasibility and accuracy of the above theoretical method are verified by test, field monitoring and ABAQUS numerical simulation, and the influencing factors of the deformation zoning type of surrounding rock of the actual roadway are analyzed. The results show that the actual surrounding rock zoning type and boundary stress value determined by the deformation zoning mechanism of deep roadway surrounding rock in this paper are more accurate than the calculation results of the traditional deformation zoning theory, and the variations in the basic mechanical properties of surrounding rock and the strength of the in-situ stress field determined the number of actual deformation zones and the boundary stress values of each zone in the roadway surrounding rock. The research results can potentially aid the support designing and stability evaluation of deep roadway surrounding rock. • Actual deformation zones in deep roadway surrounding rock were analyzed. • Novel four-stage strain-softening model determines zoning type and boundary stress. • Factors influencing the type and number of deformation zones were investigated. • Actual deformation zones rely on mechanical properties and in-situ stress of rock. • Model accuracy was verified using field tests and numerical simulations. [ABSTRACT FROM AUTHOR]

In order to solve the problem of deep soft rock roadway support difficultly in Chaohua Coal Mine, numerical simulation, theoretic analysis and site observation were applied to analyze the failure reasons. The deep soft rock roadway with original support was passive support which couldn't adapt the deformation of deep soft rock roadway. Based on the analysis, secondary roof bloting with anchor and wire mesh support technology with high strength and stability was presented. The first high strength pre-stressed bloting with wire mesh support could reinforce surrounding rock in time and form high strength bearing structure with surrounding rock, the stability and load carrying capacity of the bearing structure was improved by secondary anchor supplement support. The results showed that secondary roof bloting with anchor and wire mesh support technology was applied to deep soft rock roadway, the maximum roof-to-floor convergence was 73 mm, the maximum rib-to-rib convergence was 51 mm, the maximum roof-to-floor convergence speed was 1.62 mm/d, the maximum rib-to-rib convergence speed was 1.13 mm/d, the deformation of deep soft rock roadway was controlled effectively. [ABSTRACT FROM AUTHOR]

Coal and gas outbursts seriously threaten the mining safety of deep coal mines. The evaluation of the risk grade of these events can effectively prevent the occurrence of safety accidents in deep coal mines. Characterized as a high-dimensional, nonlinear, and small-sample problem, a risk evaluation method for deep coal and gas outbursts based on an improved quantum particle swarm optimization support vector machine (IQPSO-SVM) was constructed by leveraging the unique advantages of a support vector machine (SVM) in solving small-sample, high-dimension, and nonlinear problems. Improved quantum particle swarm optimization (IQPSO) is used to optimize the penalty and kernel function parameters of SVM, which can solve the optimal local risk and premature convergence problems of particle swarm optimization (PSO) and quantum particle swarm optimization (QPSO) in the training process. The proposed algorithm can also balance the relationship between the global search and local search in the algorithm design to improve the parallelism, stability, robustness, global optimum, and model generalization ability of data fitting. The experimental results prove that, compared with the test results of the standard SVM, particle swarm optimization support vector machine (PSO-SVM), and quantum particle swarm optimization support vector machine (QPSO-SVM) models, IQPSO-SVM significantly improves the risk assessment accuracy of coal and gas outbursts in deep coal mines. Therefore, this study provides a new idea for the prevention of deep coal and gas outburst accidents based on risk prediction and also provides an essential reference for the scientific evaluation of other high-dimensional and nonlinear problems in other fields. This study can also provide a theoretical basis for preventing coal and gas outburst accidents in deep coal mines and help coal mining enterprises improve their safety management ability.

Effective risk prevention and management in deep coal mines can reduce the occurrences of outburst accidents and casualties. To address the low accuracy and inefficiency of coal-gas outburst prediction in deep coal mines, this study proposes a deep coal-gas outburst risk prediction method based on kernal principal component analysis (KPCA) and an improved extreme learning machine (SAPSO-ELM) algorithm. Firstly, high-dimensional nonlinear raw data were processed by KPCA. Secondly, the extracted sequence of outburst-causing indicator principal components were used as the input variables for the simulated annealing particle swarm algorithm (SAPSO), which was proposed to optimize the input layer weights and implied layer thresholds of the ELM. Finally, a coal and gas outburst risk prediction model for a deep coal mine based on the SAPSO-ELM algorithm was developed. The research results show that, compared with the ELM and PSO-ELM algorithms, the SAPSO-ELM optimization algorithm significantly improved the accuracy of risk prediction for coal-gas outbursts in deep coal mines, and the accuracy rate was as high as 100%. This study enriches the theory and methods of safety management in deep coal mines, and effectively helps coal mine enterprises in improving their ability to manage coal-gas outburst risks.

The unsafe behavior of miners seriously affects the safety of deep mining. A comprehensive evaluation of miners' unsafe behavior in deep coal mines can prevent coal mine accidents. This study combines HFACS-CM, SEM, and SD models to evaluate miners' unsafe behaviors in deep coal mining. First, the HFACS-CM model identifies the risk factors affecting miners' unsafe behavior in deep coal mines. Second, SEM was used to analyze the interaction between risk factors and miners' unsafe behavior. Finally, the SD model was used to simulate the sensitivity of each risk factor to miners' unsafe behavior to explore the best prevention and control strategies for unsafe behavior. The results showed that (1) environmental factors, organizational influence, unsafe supervision, and unsafe state of miners are the four main risk factors affecting the unsafe behavior of miners in deep coal mines. Among them, the unsafe state of miners is the most critical risk factor. (2) Environmental factors, organizational influence, unsafe supervision, and the unsafe state of miners have both direct and indirect impacts on unsafe behaviors, and their immediate effects are far more significant than their indirect influence. (3) Environmental factors, organizational influence, and unsafe supervision positively impact miners' unsafe behavior through the mediating effect of miners' unsafe states. (4) Mental state, physiological state, business abilities, resource management, and organizational climate were the top five risk factors affecting miners' unsafe behaviors. Taking measures to improve the adverse environmental factors, strengthening the organization's supervision and management, and improving the unsafe state of miners can effectively reduce the risk of miners' unsafe behavior in deep coal mines. This study provides a new idea and method for preventing and controlling the unsafe behavior of miners in deep coal mines.

Affected by high ground stress, the surrounding rock control of a roadway is one of the most important factors restricting the utilization of deep resources. Therefore, it is necessary to propose a method to improve the stress environment of the deep-buried roadway and reduce its deformation. The article focuses on the 121,302 machine roadway in Kouzidong coal mine to analyze the large deformations of roadways near goafs (RNGs) in deep coal mines and reveal the mechanisms related to pressure relief via roof cutting. Through physical simulation, overburdened structures and the migration laws of RNGs in deep coal mines are studied, and the overburdened RNGs will eventually have a double short-arm "F"-type suspended roof structure. The superposition movement of the structure is the prime cause for the large deformation of the RNGs considered here. Artificial roof cutting can weaken the superposition effect of the double "F" structure and induce the roof to produce a new fracture. Meanwhile, sliding deformation along the fault line releases greater stress, and the cut roof can better fill the goaf. The stress distribution ratio between goafs and the coal pillar is improved. Here, a mechanical model of key block B' (KBB') is considered and the stability criterion of KBB' is obtained. According to the theoretical calculation here, the stress of a coal pillar could be reduced by 19.14% when KBB' is cut along the edge of the coal pillar in the 121,302 machine roadway. After engineering verification, the field observation result shows that the deformation of the 121,302 machine roadway is reduced by more than 50% after roof cutting. [ABSTRACT FROM AUTHOR]

The friction wear and thermal fatigue cracking of the brake shoe and friction-induced self-excited vibration (frictional flutter) of the disc brake can easily occur during emergency braking of a deep coal mine hoist with at high speed and with a heavy load. Therefore, tribo-brake characteristics between the brake disc and brake shoe during emergency braking of a deep coal mine hoist are investigated in the present study. Scaled parameters of the disc brake of a deep coal mine hoist are determined by employing the similarity principle. Friction tests between friction disc and brake shoe are carried out to obtain the coefficient of friction in the case of high speed and large specific pressure between the friction disc and brake shoe. Coupled thermo-mechanical finite element analyses of the brake disc and brake shoe are established to investigate temperature and stress fields of the brake disc and brake shoe during emergency braking, which is validated by the engineering failure case. Effects of braking parameters on flutter characteristics between the brake disc and brake shoe are explored by employing a double-degrees-of-freedom vibration mechanism model. The results show that the maximum temperature, equivalent Von Mises stress and contact pressure are all located at the average friction radii of contact surfaces of the brake disc and brake shoe during emergency braking. The cage crashing accident in the case of high speed and heavy load in a typical coal mine shows crack marks and discontinuous burn marks at central locations of brake shoe and brake disc surfaces, respectively, which indicates frictional flutter characteristics between brake disc and brake shoe. During emergency braking, flutter time duration decreases with increasing initial braking speed and damping parameter; the flutter amplitude and frequency of the disc brake increases with increasing normal braking load and stiffness, respectively. [ABSTRACT FROM AUTHOR]