Soft and hard composite rock strata are frequently encountered in transportation, geotechnical, and underground engineering. However, most of the current support is designed for homogeneous rock masses, which ignores the different anchoring effect in soft and hard composite rock strata. A numerical study is presented in this paper on the pull‐out behavior of fully grouted rock bolts in soft and hard composite rock strata. The nonlinear bond‐slip relationship of bolt‐grout interface that is anchored in soft rock and hard rock is obtained from laboratory test, respectively. Then, the nonlinear bond‐slip relationship is put into the numerical model. The numerical result shows a close match with the experiment tests and the proposed model. Lithological sequence, layer thickness ratio, and layer numbers are taken into consideration in numerical simulation models. Under the same layer number, the shallower‐soft and deeper‐hard composite rock strata (SHCRS) have a higher bearing capacity and deformation resistance than the shallower‐hard and deeper‐soft composite rock strata (HSCRS). As the soft‐to‐hard thickness ratio in SHCRS increases, the initial stiffness of the load‐displacement curve and peak load decreases continuously. The load‐displacement curve shows the same initial stiffness for different hard to soft thickness ratios in HSCRS. As the hard to soft thickness ratio increases, the load peak and the displacement at the peak load increase. Therefore, the closer the hard rock is to the loading end, the greater the initial stiffness of the load‐displacement curve is. The greater the hard rock thickness, the larger the peak load. Under the same anchor length, the peak load and the displacement at the peak load decrease with the increase of layer numbers, but the reduction magnitude decreases. This paper leads to a better understanding of the load transfer mechanism for the anchoring system in soft and hard composite strata and provides a reference for scientific support design and evaluation method.

Large deformation of roadway and coal bump failures have always been the focus in deep underground engineering. By considering the Lu’an mining district in China, the failure mode and stability improvement process of high‐stress roadways were analysed with the field tests and numerical simulations. The field test results showed that a great amount of deformation and serious damage occurred in surrounding rocks during panel retreat due to the suspended roof. A novel approach employing roof fracturing and collapsed rock filling effect was adopted to maintain the roadway stability. A numerical model was established with the Universal Distinct Element Code (UDEC) to research the fracturing characteristics between the roadway and gob roofs and the stress change in the surrounding rock. The modelling results demonstrated that, without fracturing roof, the peak vertical stress of the coal pillar was 18.3 MPa and the peak vertical stress of the virgin coal rib was 15.6 MPa. The roadway was in a state of high stress. With fracturing roof, the peak vertical stress of coal pillar was 9.3 MPa and the peak vertical stress of virgin coal rib was 13.4 MPa. The fractured rock mass in the gob expanded in volume and provided supporting resistance to the overlying strata, which relieved stress concentrations in the coal pillar. Field measurement results indicated that the roadway large deformation was successfully resolved during excavation and panel retreat after implementing the novel approach, providing useful references for the application of this novel approach in similar coal mines.

A coal-rock system is a common combination form in coal mines. In order to explore the energy interchange law of a coal-rock combined body and the interaction relationship between the two bodies, loading tests of coal-rock combined bodies with different height ratios were carried out. The loading path of rock in coal-rock combined bodies was demonstrated by means of a single loading and unloading test of the same-sized rock sample. Furthermore, a method to calculate rock energy was proposed based on the area of the loading and unloading curve. The experimental results show that the greater the surrounding rock pressure is, the smaller increase rate of lateral and volumetric strain in the postpeak stage will be when the same height ratio is present. An increase in the surrounding rock pressure causes an increase in the total strain energy density of small-sized rock samples. However, the total strain energy density is always greater than the elastic strain energy density. And the elastic and dissipated strain energy densities also increase, along with the energy dissipation with unloading. When the proportion of coal bodies increases, the energy accumulation also shows an increasing trend. When the height of the coal is greater than half the height of the complete specimen, the coal energy proportion is greater than 60%. After reaching the yield load, the energy in the coal body is dissipated in forms such as plastic deformation, internal damage, block friction, radiation energy, and kinetic energy. Therefore, the energy released is, in part, reflected in the rock body.

Roadway deformation and coal bumps are major challenges for underground engineering. Taking the Lu'an mining district in Shanxi Province as an example, the failure mechanism of deep high-stress roadway was studied. Numerical simulations and the similar material test were performed to investigate the effect of the roof presplitting and rock mass filling on the stability of roadway surrounding rock. The change laws and distribution features of stress and plastic zone, and the fracture characteristics of roof strata were analyzed. Under conditions where the roof was not presplit, the roadway was in a high-stress environment due to the massive suspended roof. The failure process of the high-stress roadway began with tensile crack in the shallow roof and then extended to the two ribs, resulting in the yield of a coal pillar. Under the condition that the roof was presplit, the peak value of the vertical stress at the coal pillar was decreased from 18.2 to 9.8 MPa while that of the virgin coal rib was decreased from 15.8 to 13.5 MPa. Both the similar material test and numerical results showed that the Roof outside the Coal Pillar (RCP) was cut off along the presplitting lines and was fractured into different sizes of masses. The broken rock mass filled in the goaf and supported on the overlying strata, which reduced the load on the coal pillar. The field monitoring data indicated that implementing the roof presplitting and rock mass filling processes successfully controlled large deformations of the roadway and coal bumps, which improved the stability of surrounding rock.

Severe dynamic disturbance in extrathick coal seam mining has become one of the main factors threatening the stability of roadway surrounding rock. In this article, the #6 thick coal seam of Buliangou mine in Inner Mongolia, China, is taken as the engineering background. A mechanical model of the roadway roof structure is established to obtain an analytic formula of the key block subsidence. A three‐dimensional discrete element model is established and used to verify the field measurement results. The fracture characteristics of the main roof above the F6104 transport roadway and the deformation and damage evolution law of the surrounding rock during thick coal seam mining are analyzed. The results show that because of the long‐term breaking and falling of the roof rocks during extrathick coal seam mining, the F6104 transport roadway will undergo two severe mining disturbances at the locations of 10∼30 m and 50∼70 m ahead of the F6103 working face. During the two disturbance periods, the roadway roof displacement settles to 300∼350 mm and 750∼800 mm, and the deformation of the solid coal wall reaches 650∼700 mm and 1350∼1450 mm, respectively. The energy change curve of the total length of the fractured key roof is obtained, and when mining at 50 m, the basic roof is close to its tensile strength, and the strain energy can reach the peak value of 5.2 × 10⁴ kJ, which easily leads to rock burst. The plastic damage zones on both sides of the roadway develop to the roof central area and eventually coalesce, and the deformation of the surrounding rock is obvious. When mining at 50∼70 m, the basic roof breaks and unloads, and elastic strain energy of 3.57 × 10⁴ kJ is instantaneously released. These two dynamic disturbances are the main reasons for the instability of the roadway surrounding rock. The results clarify that the failure mechanism investigation of roadways in thick coal seam mining conditions can be effectively applied to control the stability of the roadway surrounding rock under strong mining disturbance.

The width design of protective pillars is an important factor affecting the stability of high‐stress roadways. In this study, a novel numerical modeling approach was developed to investigate the relationship between protective pillar width and roadway stability. With the 20 m protective pillar width adopted in the field test, large deformation of roadways and serious damage to surrounding rocks occurred. According to the case study at the Wangzhuang coal mine in China, the stress changes and energy density distribution characteristics in protective pillars with various widths were analysed by numerical simulation. The modeling results indicate that, with a 20 m wide protective pillar, the peak vertical stress and energy density in the pillar are 18.5 MPa and 563.7 kJ/m³, respectively. The phenomena of stress concentration and energy accumulation were clearly observed in the simulation results, and the roadway is in a state of high stress. Under the condition of a 10 m wide protective pillar, the peak vertical stress and energy density are shifted from the pillar to roadway virgin coal region, with peak values of 9.5 MPa and 208.3 kJ/m³, respectively. The decrease in vertical stress and energy density improves the stability of the protective pillar, resulting in the roadway being in a state of low stress. Field monitoring suggested that the proposed 10 m protective pillar width can effectively control the large deformation of the surrounding rock and reduce coal bump risk. The novel numerical modeling approach and design principle of protective pillars presented in this paper can provide useful references for application in similar coal mines.