The Stability Factors’ Sensitivity Analysis of Key Rock B and Its Engineering Application of Gob-Side Entry Driving in Fully- Mechanized Caving Faces

To reveal the critical factors of the main roof influencing stability of surrounding rocks of roadways driven along goaf in fullymechanized top-coal caving faces, this paper builds a structural mechanics model for the surrounding rocks based on geological conditions of the 8105 fully-mechanized caving face of Yanjiahe Coal Mine, and the stress and equilibrium conditions of the key rock block B are analyzed, and focus is on analyzing rules of the key rock block B influencing stability of roadways driven along goaf. +en, the orthogonal experiment and the range method are used to confirm the sensitivity influencing factors in numerical simulation, which are the basic main roof height and the fracture location, the length of the key rock block B, and the main roof hardness in turn. It is revealed that the basic main roof height and its fracture location have a greater influence on stability of godside entry driving. On the one hand, the coal wall and the roof of roadways driven along goaf are damaged, and the deformation of surrounding rocks of roadways and the vertical stress of narrow coal pillars tend to stabilize along with the increase of the basic main roof height. On the other hand, when the gob-side entry is located below the fracture line of the main roof, the damage caused by gob-side entry is themost serious.+erefore, on-site gob-side entry driving should avoid being below the fracture line of the main roof. At last, industrial tests are successfully conducted in the fully-mechanized top-coal caving faces, 8105 and 8215, of Yanjiahe Coal Mine.


Introduction
Adoption of the narrow coal pillar gob-side entry driving technique can efficiently increase recycling efficiency of coal resources and control effect of surrounding rocks of roadways. e technique has been widely used in China, but stability of surrounding rocks of roadways driven along goaf has limited further development of the technique [1][2][3][4][5][6][7][8][9][10][11][12][13][14][15][16][17]. Research has shown that there are multiple factors that influence the stability. References [1][2][3][4][5][6][7][8][9][10][11][12][13][14][15][16][17] are the Chinese scholars' findings. Among them, Bai et al. [1,2] built an arcshaped triangular block structural mechanism model for surrounding rocks based on stress conditions of the working face; Li [3,4] analyzed influence of big and small structures of roadways driven along goaf in fully-mechanized top-coal caving face on stability of surrounding rocks; Zhang et al. [5,6] analyzed deformation features of surrounding rocks of gob-side entry retaining in the fully-mechanized coal face with top-coal caving and influence of multiple factors on deformation and stress of gob-side entry retaining with entry-in packing in the top-coal mining face; Wang et al. [7,8] discussed influence of three different fracture locations of the overlaying main roof of roadways driven along goaf on stability of surrounding rocks of the roadway and rational width of the narrow coal pillar based on the fracture line location of key rock B in the main roof; Zhu et al. [9] studied the main factors affecting the deformation of the filling body; Xie [10] adopted UDEC simulation to analyze rules of six factors, such as support technique and influencing stability of surrounding rocks of gob-side entry retaining in top-coal mining faces; Zhang et al. [11,12] analyzed four different Chinese coal mining sites and evaluated the influencing factors; Meng and Li [13] studied parameters' sensitivity of bolt support in gob-side entry driving and found that the main influence parameters were bolt pretightening force and bolt interval; Yuan et al. [14] studied the dynamic effect and control mechanism of key strata in the immediate roof; Bai et al. [15] studied the stress state and the deformation failure mechanism of the heading adjacent to the advancing working face roof structure; Li et al. [16] studied the in situ stress distribution characteristics of the rock mass near different slope angles' hillslope surfaces; Wang et al. [17] studied the stability of the concrete artificial side introduced to stabilize a gob-side entry.
e above analysis suggested that stability of the key rock block B could directly influence stability of the large structure of roadways driven along goaf and indirectly influence the stress environment of the small structure. e former mainly adopted theoretical analysis and quantitative experiments to analyze stability of the large structure of roadways driven along goaf, while the latter mainly analyzed influence of factors, such as support parameters, on stability of surrounding rocks of roadways based on numerical simulation. Few of them analyzed sensitivity of major factors of the main roof on stability of surrounding rocks of roadways driven along goaf in fully-mechanized top-coal caving faces. In view of the research gap, this paper built a structural mechanics structure for surrounding rocks of the key rock block B to quantitatively analyze influence of main roof factors on stability of surrounding rocks. Based on mining technical conditions of Yanjia Coal Mine in Binchang Mining Area, four factors of the main roof are confirmed. Combining the orthogonal experimental scheme, the author analyzes sensitivity of the four main roof factors and conducts a univariate analysis of the two factors with a stronger sensitivity. Research findings are successfully applied to the 8105 working face and further promoted in the 8215 working face. erefore, research results of this paper can contribute to popularization of gob-side entry driving technique under similar geological conditions.

Building of the Key Rock Block B Mechanical Model.
Currently, the narrow coal pillar gob-side entry driving technique is prevailing in China. After the overlaying rock of the goaf in the working face of the former sector stabilizes, the roadway driven along goaf starts. Figure 1 shows the structure of the fracture line of the main roof surrounding rock structure within the coal wall during gob-side entry in a fully-mechanized top-coal caving face.
Below is the stress analysis of the key rock block B [18]: the resultant force of shear force and horizontal thrust of the key rock block A on the key rock block B is R AB and T AB , respectively; the resultant force of vertical shear force and horizontal thrust of the key rock block C on the key rock block B is R BC and T BC , respectively; the dead load of the soft stratum above the block B is F R ; the dead load resultant force of the block B is F Z ; the support force of the goaf waste rock for the block B is F G ; the support force of the damaged immediate roof in the section without top-coal caving for the block B is F D ; the support force of the immediate roof in the narrow coal pillar for the block B is F S ; the support force of the immediate roof of the physical coal wall for the block B is F M ; the rotation angle of the block B is θ. e stress situations of block B are shown in Figure 2. To put the surrounding rock structure of the block B in an equilibrium state, the block B should achieve equilibrium of stress horizontally and vertically after its stability. Besides, there is no allowance for slipping and rotation.
(3) Conditions for the key rock blocks B and A to lose stability are as follows [19]: where tanφ stands for the friction factor between the two rock blocks, (T AB /L 1 a) stands for the average crushing stress between the two rock blocks, MPa, η stands for the special coefficient of the stress analysis between the two rock blocks, and σ C stands for the compressive strength of the rock block B, MPa. When the block B contacts the two and generates compression, the latter will provide support stress for the main roof. e value of the stress is closely related to the compression rate of the immediate roof and the goaf waste rocks, and the compression rate has a close bearing on thickness, strength, and swell factor of the immediate roof and the coal strata and the mechanical strength of goaf waste rocks. When the fracture line location of the main roof is different [8], the horizontal stress remains the same, but the vertical stress of the key rock block B changes. See the following:

Analysis of Factors
(1) When the fracture line of the main roof is within the physical coal wall, Under the condition, the vertical stress changes as follows: (2) When the fracture line of the main roof is outside the narrow coal pillar, Under the condition, there is no F M on the block B vertically, and the stress changes as follows: (3) When the fracture line of the main roof is just above the roadway, Under the condition, there is no F M and F S on the block B vertically, and the stress changes as follows: To sum up, different fracture line locations of the main roof will directly influence the value of the vertical stress on the block B. Besides, during the rotary subsidence process, different compression ratios with the immediate roof can result in different value of support stress. erefore, there are many factors influencing stability of the block B. It is hard to accurately and quantitatively describe its value theoretically and in practices. It is necessary to turn to the numerical simulation to analyze influence of factors of the basic roof on stability of surrounding rocks in gob-side entry in fullymechanized top-coal caving faces.

Confirmation of Influencing Factors.
e working face of Yanjiahe Coal Mine features a single-wing main roadway belt layout. In order to efficiently connect working faces and efficiently control deformation of the surrounding rock of roadways, the narrow coal pillar gob-side entry is adopted, and favorable technical and economic efficacy is achieved [20][21][22][23], but geological conditions of Yanjiahe Coal Mine are complex, and the working face mine pressure behaviors and drill peeping analysis show that storage condition changes of the main roof are huge [2,3]. Based on geological conditions of Yanjiahe Coal Mine, rules of four factors, including the main roof height and strength, the length of the key rock block B, and the fracture location of the main roof, in influencing stability of surrounding rocks of roadways driven along goaf is analyzed.

Orthogonal Experiment.
e orthogonal experimental scheme is adopted to analyze different degrees of influence of four factors of the main roof on stability of surrounding rocks of roadways driven along goaf. e experimental scheme and the four factors are shown in Tables 1 and 2. Under the same bolt support, the deformation of surrounding rocks of roadways is an index to measure influence of different factors on stability of surrounding rocks of roadways. According to the orthogonal experimental scheme, the sensitivity of the four factors is ranked in order, and the two of them with a greater sensitivity are selected for the univariate analysis as to their influence on stability of surrounding rocks of roadways.

Modeling.
Based on geological conditions of the 8105 fully-mechanized top-coal caving face, the authors adopt UDEC2D4.0 to build the numeric calculation model [21,24,25]. e constitutive relation between the rock block and the surrounding rocks features a Mohr-Coulomb model. e physical mechanical parameters of the rock mass are shown in Table 3. e belt cross-heading fracture surface of 8105 is 4000 mm × 3000 mm; and, the combined support featuring high-intensity deformed steel bar resin bolt + high preload + cable reinforcement + roof bolting with bar and wire [18,22] is adopted. During the simulation process, the neighboring caving faces are first excavated and balanced; and, then comes the gob-side entry driving. e whole process should lay out corresponding stress and displacement monitoring points.  Figure 1: Surrounding rock structural mechanics model during gob-side entry in a fully-mechanized top-coal caving face.

Analysis of the Orthogonal Experimental Scheme.
e "cross-measurement method" is adopted to measure deformation of surrounding rocks. en, the range method is used to comparatively analyze the deformation of four factors under three levels. e deformation of surrounding rocks is shown in Table 2.
From Table 2, it can be seen that the range of the four factors influencing overall stability of surrounding rocks of roadways, including the main roof height and hardness, the length of the key rock block B, and the fracture location, based on the orthogonal univariate analysis is 393.54 mm, 29.19 mm, 84.20 mm, and 93.03 mm, respectively; the range of the four factors influencing stability of surrounding rocks on two ribs of roadways is 173.09 mm, 54.59 mm, 3.96 mm, and 29.02 mm, respectively; and, the range of four factors influencing stability of the roof-to-floor surrounding rocks is 220.45 mm, 25.40 mm, 80.24 mm, and 90.18 mm, respectively. us, the degree of influence of the four factors of the main roof on stability of surrounding rocks of roadways driven along goaf can be ranked as follows: the main roof height ＞ the main roof fracture location ＞ the length of the key rock block B ＞ the main roof hardness. e degree of influence of the four factors on surrounding rocks on two ribs of roadways can be ranked as follows: the main roof height ＞ the main roof hardness ＞ the main roof fracture location ＞ the length of the key rock block B. e degree of influence of the four factors on the roof-to-floor surrounding rocks can be ranked as follows: the main roof height ＞ the main roof fracture location ＞ the length of the key rock block B ＞ the main roof hardness. It can be seen that the main roof height and the main roof fracture location are two factors with a greater sensitivity.

Simulation Scheme.
In order to further analyze the influence of the main roof height and its fracture location on stability of roadways driven along goaf, the univariate analysis is adopted. e simulation scheme is shown in Table 4. e main roof hardness is set to be medium-hard, and the length of the key rock block B is set to be 18 m.

Influence of the Main Roof Height on Stability of Surrounding Rocks of Roadways.
e average value of the deformation of surrounding rocks of roadways and the vertical stress value before and after coal pillar driving of nine schemes included in Table 2 are shown in Figure 3.
From Figure 3, it can be seen that the coal pillar stress before driving is larger than that after driving. Before the roadway driving, the overlaying strata of neighboring caving faces is basically stable, the key rock block B is in an equilibrium state, and the stress on the coal wall is concentrated. In Schemes 1-3, Schemes 4-6, and Schemes 7-9, the vertical stress scope in the central position of the narrow coal pillar is 7.41∼9.43 MPa. is suggests that the influence of the main roof height and the fracture location on the coal pillar stress before the roadway driving is not significant. After the roadway driving, the stress of surrounding rocks of roadways needs to be redistributed, and the stress peak value transfers to the inside of the coal. At the moment, the coal pillar is in the stress declining area. In Schemes 1-3, Schemes 4-6, and Schemes 7-9, the average vertical stress of the coal pillar is within the range of 4.36∼5.03 MPa, 1.56∼2.67 MPa, and 2.07∼2.38 MPa, respectively. is suggests that, when the main roof height is 0 m, the narrow coal pillar has a greater support stress, F S , for the key rock block B. Along with the increase of the main roof height, the support stress of the main roof in the section without coal and the goaf waste rocks for the key rock block B is F D and F G , respectively, thus reducing the compression of the key rock block B on the immediate roof above the coal pillar.
From Figure 3, it can be seen that the relative deformation on the two ribs of the roadway in Schemes 1-9 is 151.50%, 141.45%, 136.04%, 163.12%, 143.72%, 131.24%, 168.07%, 76.96%, and 146.21% of the roof-to-floor relative deformation. us, it can be judged that deformation on two ribs dominates in roadways driven along goaf in fullymechanized top-coal caving faces. e relative deformation on two ribs in Schemes 1-3, Schemes 4-6, and Schemes 7-9 is 274.95∼322.25 mm, 315.01∼368.86 mm, and 295.57 mm∼390.31 mm, respectively; the deformation is small in Schemes 1-3. e deformation in Schemes 4-6 and Schemes 7-9 is almost the same. After roadways driven along goaf, the surrounding rocks of roadways obtain timely support. e surrounding rocks' stress is quickly redistributed. Relying on bolt and cable support, surrounding rocks of roadways form a small structure. With the rotary subsidence of the key rock block B, the support stress provided by the immediate roof and the physical coal wall right above the narrow coal pillar, F S and F M , increases and the stress of the narrow coal pillar and the physical coal wall concentrates, thus resulting in expansion and deformation of surrounding rocks.

Influence of the Main Roof Fracture Location on Stability of Surrounding Rocks of Roadways.
e deformation of surrounding rocks of roadways in Schemes 1∼3,

Advances in Civil Engineering
Schemes 4∼6, and Schemes 7∼9 is shown in Figures 4(a)-4(c), respectively. From Figure 4, it can be seen that deformation of surrounding rocks of roadways driven along goaf mainly happens in the physical coal wall and the roof side. Under the same height but different basic roof fracture locations, the deformation of surrounding rocks of roadways is different. Below is a detailed analysis. At the moment, there is no immediate roof above the coal strata, thus resulting in no goaf waste rock support force, F G , for the key rock block B. When the basic main roof fracture location is right above roadways, the key rock block B will rotate and sink directly on the roof side of the roadway and the side of the narrow coal pillar. Under the condition, the deformation of surrounding rocks of roadways is larger than that under the other two conditions.
When the main roof height is 5 m, the deformation of the physical coal wall side, the roof side, the coal pillar side, and the floor side is 360. 9     Advances in Civil Engineering pillar, the roadway roof, the physical coal wall, and the goaf to achieve an equilibrium state. When the main roof fracture location is right above roadways, the compression rate between the narrow coal pillar and the immediate roof right above roadways is relatively large, and the stress concentration is high. Consequently, the deformation of surrounding rocks of the narrow coal pillar side and the roof side becomes larger than that under the above two conditions. erefore, in order to reduce the stress concentration degree of the narrow coal pillar side and right above roadways, it is necessary to avoid laying gob-side entry driving right below the main roof fracture location. In this way, the bolt or the cable will control deformation of surrounding rocks of roadways more easily.

Engineering Application Analysis
Based on the above research, engineering application analysis is conducted on two fully-mechanized top-coal caving faces, namely, 8105 and 8215. erefore, the coal pillar width should not be within the range of 3.42∼3.87 m [8,22]. (See Figure 5) e numerical simulation analysis is adopted to analyze changing rules of surrounding rock deformation and coal pillar and seam floor vertical stress under different coal pillar width. Finally, the reasonable width of the coal pillar is 6.5 m. After 20 days of roadway driving, the surrounding rocks are basically stable, and the maximum roof-to-floor and two-rib displacement is 118 mm and 65 mm, respectively. Influenced by the caving face excavation, the maximum roof-to-floor and two-rib displacement is 200∼420 mm and 380∼600 mm, respectively. e control effect of surrounding rocks of roadways is shown in Figure 6.

8215 Fully-Mechanized Top-Coal Caving Face.
e 8215 track transport roadway drives along the 8214 goaf boundary, and the coal pillar width of the coal section is 8.0 m. Six monitoring stations are laid out to monitor surrounding rocks of roadways. After 20 days of roadway driving, the surrounding rocks are basically stable, the bolt and cable working conditions are favorable, and the maximum roof-floor and two-rib displacement amount of six monitoring stations is 100∼160 mm and 80∼100 mm, respectively. Influenced by excavation, the maximum rooffloor and two-rib displacement amount is 220∼460 mm and 340∼620 mm, respectively. e stress variation rules of the coal monitored by the monitoring station 2 are shown in Figure 7. 064Y, 065Y, and 066Y in Figure 7 represent the surrounding rock stress 4 m within the narrow coal pillar, 5 m within the physical coal wall, and 10 m within the physical coal wall.
From Figure 7, it can be seen that the stress 4 m within the coal pillar is low and changes within the range of 0.1∼0.2 MPa. e stress 5 m within the physical coal wall is large. If not influenced by excavation, it changes within the range of 0.1∼0.6 MPa, and if influenced by excavation, it changes within the range of 0.7∼0.8 MPa. e stress 10 m within the physical coal wall is relatively large. If not influenced by excavation, the change is within the range of 0.1∼0.8 MPa; if influenced by excavation, the peak stress value can reach 10.9 MPa. Results suggest that, if the coal is not influenced by excavation, the surrounding rock stress is relatively small; if influenced by excavation, the surrounding rock stress is transferred to be within the physical coal wall. However, the roadways driven along goaf far always in the stress decline area, which is beneficial to the surrounding rock control of roadways.
Changing rules of the surrounding rock deformation and the coal stress suggest that roadways, influenced by excavation, are efficiently controlled, thus meeting requirements of the working face stopping [20,23]. In order to master the main roof fracture line of the 8215 caving face, probing drills are laid by the 8215 track transport roadway to the overlaying strata on the 8214 goaf, and the probing results are shown in Figure 8.
From Figure 8, the large fractures concentrate above the coal pillar near the goaf. eir distance away from the coal pillar margin is short, namely, 1     Advances in Civil Engineering right of the coal pillar in goaf and far away from the coal pillar margin. It suggests that it is reasonable to set the coal pillar to be 8.0 m, which can avoid putting the roadway right under the basic fracture line, and the surrounding rock control effects are shown in Figure 9.

Conclusions
is paper builds the structural mechanics model of the key rock block B, points out that equilibrium of the key rock block B should meet conditions not to lose equilibrium horizontally and vertically, qualitatively analyzes the influence of the main roof factors on stress of the key rock block B, and places particular emphasis on stresschanging rules of the key rock block B vertically under three different fracture locations of the main roof.
Based on mining technical conditions of Yanjiahe Coal Mine in Binchang Mining Area, four factors of the main roof are confirmed, including the main roof height and hardness, the length of the key rock block B, and the main roof fracture location.
e orthogonal experiment is used to set nine plans; the range method is employed to confirm the sensitivity order of the main roof, which is as follows: the main roof height ＞ the main roof fracture location ＞ the length of the key rock block B ＞ the main roof strength.
e univariate method is adopted to analyze the influence of the main roof height and the fracture location on stability of surrounding rocks of roads in fullymechanized caving faces. It is found that surrounding rock deformation during roadways driven along goaf happens in the physical coal wall and the roof side. e deformation of surrounding rocks of roadways and the vertical stress of the narrow coal pillar tend to stabilize along with the increase of the main roof height. Meanwhile, the roadway driving right below the basic fracture line should be avoided. e engineering application effects suggest that the narrow coal pillar layout is reasonable on the 8105 and 8215 caving face, which can avoid putting the main roof fracture location right above the roadway. e surrounding rocks of the driving roadways are in a low-stress environment, and the surrounding rocks of the roadways are efficiently controlled. All these have guaranteed safe and efficient stopping of caving faces.

Data Availability
e data used to support the findings of this study are included within the article.