3.1.2. The Effects of Working Face Mining on the Roadway Surrounding Rock

The 42202 transport roadway was affected by the stress transfer from the floor of the coal pillar in the overlying coal seam and the lateral abutment pressure generated in the stope mining process of the adjacent 42201 working face, respectively. The structural evolution of the overlying roof rock of the 42201 working face is shown in Figure 6a. After the coal seam is mined, the overlying rock structure of the gob roof fails, and the stress state is changed significantly. With the continuous advancement of the working face, the scope of the gob increases, and the immediate roof fracture and expand to fill the gob. As "O-X" periodic fracture occurs in the main roof, block B forms an arc triangle with the sectional coal pillar. One end of rock block B rotates and contacts the gangue in the gob, and the other end breaks off inside the rib of the sectional coal pillar. Despite the rotation and sinking, rock block B interlocks with rock block C and rock mass A, forming a hinge "macrostructure" [25,26]. In this process, the static force system transferred by the overlying remaining coal pillar is *σ*<sup>1</sup> <sup>1</sup> ; The static force system generated by roadway surrounding rock excavation disturbance is *σ*<sup>2</sup> <sup>1</sup> ; The lateral abutment pressure system due to mining at the 42201 working face is *σ*<sup>3</sup> <sup>1</sup> . Therefore, the total stress on the roadway surrounding rock *σ* can be expressed as *σ* = *σ*<sup>1</sup> <sup>1</sup> + *<sup>σ</sup>*<sup>2</sup> <sup>1</sup> + *<sup>σ</sup>*<sup>3</sup> <sup>1</sup> , as shown in Figure 6.

A micro-element from the roadway rib shown in Figure 6b is selected for analysis. The surrounding rock stress states before and after multiple disturbances and the effect on its stability are illustrated in Figure 7.

(1) Before the disturbance from coal seam mining, the original rock stress is distributed in the rock seam. According to field measurements [27], the ratio of the maximum principal stress *σ*<sup>0</sup> <sup>1</sup> to the minimum principal stress *<sup>σ</sup>*<sup>0</sup> <sup>3</sup> is approximately 1.3, i.e., the difference between the tangential stress and the radial stress is small. Thus, the diameter of the Mohr's circle (*σ*<sup>0</sup> <sup>1</sup> − *<sup>σ</sup>*<sup>0</sup> <sup>3</sup> ) representing the stress state is very small and below the strength envelope *L*1.

(2) As the overlying coal seam (2-2 coal seam) is mined, the tangential stress in the rock mass increases from *σ*<sup>0</sup> <sup>1</sup> to *<sup>σ</sup>*<sup>1</sup> <sup>1</sup> due to the stress transfer of the remaining coal pillar abutment pressure, and the underlying coal seam enters a high pressure-bearing state.

(3) At the moment of excavation and unloading of the 42202 stope mining roadway, the radial stress around the perimeter of the roadway decreases from *σ*<sup>0</sup> <sup>3</sup> to *<sup>σ</sup>*<sup>1</sup> <sup>3</sup> and decreases to zero at the roadway surface. However, the tangential stress in the surrounding rock increases rapidly from *σ*<sup>1</sup> <sup>1</sup> to *<sup>σ</sup>*<sup>2</sup> <sup>1</sup> . Thus, the radius of the Mohr's circle (*σ*<sup>2</sup> <sup>1</sup> − *<sup>σ</sup>*<sup>1</sup> <sup>3</sup> ) increases significantly, and the Mohr's circle exceeds the strength envelope *L*1. It is thus evident that great deviatoric stress is generated in the surrounding rock after the roadway excavation, resulting in plastic failure of the surrounding rock.

(4) After the stope mining at the 42201 working face, the key stratum of the roof rock shows rotary movements. Due to the superposition of lateral abutment pressure transfer, the tangential stress of the surrounding rock increases sharply from *σ*<sup>2</sup> <sup>1</sup> to *<sup>σ</sup>*<sup>3</sup> <sup>1</sup> , and the plastic failure of the surrounding rock is intensified. In addition, after the large-scale compression-shear failure of the surrounding rock, dislocation failure occurs in the failure area along the shear surface. The strength of the rock mass decreases significantly in the plastic area, and the bearing capacity decreases. At this point, the rock strength envelope develops from line *L*<sup>1</sup> to line *L*2.

**Figure 6.** Roof fracture structure after mining at the working face. (**a**) Mining at the 42201 working face; (**b**) Gob-side entry retaining masonry beam structure.

3.1.3. Reinforcement Support for Stope Mining Roadway

The original support of the 42202 stope mining roadway mainly include bolts and anchor cables. According to the above numerical simulation results, when the coal pillar width is between 25 and 30 m, the plastic area in the roof is 2.5 m, and that in the two ribs is 3.0 to 3.5 m. Supported with anchor cables 6.3 m long, the bearing capacity is sufficient for the deep surrounding rock. Thus, the surrounding rock of the roof has good stability. The two ribs are only supported by bolts 2.1 m long, and the full length of the bolts is within the plastic area. As a result, the surrounding rock of the supported section has a relatively small confining pressure, and the support capacity of the bolts is greatly reduced, rendering the surrounding rock prone to overall instability. Therefore, in addition to leaving reasonable coal pillars and pressure relief, reinforcement support must be applied to the surrounding rocks of the two ribs to improve their stability and bearing capacity.

The following key issues should be noted when applying reinforcement support [28]. (1) The anchorage blind area of the weak coal mass must be controlled. The anchorage blind area is in the shallow triangle between the compression areas (Figure 8), and the deformation and loosening of the coal mass in this area weaken the bolt support force, forming a vicious circle. Appropriate increase in bolt density and the use of bolt and metal strip combination support are the main methods to control the anchorage blind areas. Higher bolt density reduces the range of the anchorage blind area, while steel strips spread the bolt supporting force to the shallow rock mass, thus restraining the deformation of the anchorage blind area.

**Figure 8.** Supporting effect of anchor cable and anchor belt on broken surrounding rock.

(2) The bolt-supported area has low load-bearing capacity. Considering the relatively large range of the plastic area in the surrounding rock, it is difficult for the bolts to penetrate deeply into the stable rock mass. At this time, the main function of the bolts is to form the loose surrounding rock into a supporting structure with sound stability. However, long-term stability requires combining long and short bolt support, i.e., establishing a foundation bearing ring with short bolts in the shallow part and a thick reinforcing ring in the deep part with long anchor cables. Thus, the reinforced bolt-supported area is linked to the deep stable rock, and the range of the surrounding rock-bearing ring is expanded. Consequently, fracture interconnection and loosening fragmentation failures in the shallow surface are prevented, and a deep bolt layer is formed to resist disturbance, as shown in Figure 9.

**Figure 9.** Surrounding rock-bearing structure after reinforcement and support with anchor cable.

#### *3.2. Numerical Simulation*

Based on the actual situation of the working face and the roadway at the site, a 770 m × 600 m × 175 m numerical model was established using the finite difference software FLAC3D, as shown in Figure 10. The vertical movement is restricted at the bottom of the model, and the horizontal movement is restricted at the front, rear, and both ends. The overlying rock unit weight is replaced with a uniform load of 9.6 MPa applied on the top of the model, and the lateral pressure coefficient is set to 1.3 [27]. The coal seam was simulated with the Mohr-Coulomb model, and the initial support components of the roadway were simulated using FLAC3D built-in structural units.

**Figure 10.** Global 3D numerical model.

The simulation scheme and process are as follows. (1) Application of initial ground stress and initial equilibration. (2) Progressive excavation of the 22201 and 22202 working faces. (3) In order to investigate the effects of the overlying remaining coal pillar and the adjacent 42201 working face on the 42202 stope mining roadway, the 42202 roadway was excavated when the coal pillar width was 10, 15, 20, 25, and 30 m and immediately supported after excavation. (4) The 42201 working face was gradually excavated after equilibrium calculation until the coal seam excavation was completed.

### **4. Results**

#### *4.1. Results of Theoretical Analysis*

Taking the location of the 42202 stope mining roadway below the T-shaped remaining coal pillar as an example, the vertical stress distribution directly below the T-shaped remaining coal pillar (inclined to the profile along the working face) calculated based on the stress distribution model in Figure 5 and Equations (2) and (3) is shown in Figure 11, and the vertical stress can be obtained as approximately 13.45 MPa. Therefore, the location of the roadway is affected by the abutment pressure transfer of the overlying coal pillar, and the vertical stress in the surrounding rock has reached 1.4 times the original rock stress before roadway excavation in the underlying coal seam.

In addition to effectively improving the roadway stability and reducing the surrounding rock maintenance cost, the reasonable coal pillar size also greatly reduces the waste of coal resources. The surrounding rock of reusable roadways is generally affected by the superimposed disturbance from roadway excavation and stope mining at the adjacent working faces, especially with roadways below remaining coal pillars. Therefore, the sectional coal pillar size in the underlying working face should be fully considered and designed.

**Figure 11.** Abutment pressure distribution on the T-shaped remaining coal pillar.

In accordance with conventional coal pillar width determination methods, an elastic core area of 2 times the mining height should be reserved in the middle of the coal pillar during the service of the roadway. Considering the above numerical simulation results, the distribution range of abutment pressure from the remaining coal pillar and adjacent working face, a reasonable roadway should avoid the peak superimposed abutment pressure area and be arranged in the pressure-reducing section. Therefore, the coal pillar width should not fall below 20 m.
