# Study on Failure Mechanism and Control Technology of Roadway Surrounding Rock under Complex Occurrence Coal Seam

Authors: Hai Wu; Jiaren Chen; Xu Gao; Tao Ling; Xuan Zhang; Qian Jia
Journal: Geofluids
 (2022)

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## Abstract

Aiming at the characteristics of large deformation and long deformation time under the mining influence of the 2254 east floor roadway surrounding rock, which under the condition of complex coal occurrence in Jiahe Mine. Through the actual measurement, it is found that the mineral composition of the roadway surrounding rock is mainly siderite, and it also contains 9.52% of kaolinite minerals, and the surrounding rock is relatively weak. Through drilling and peeping, it is found that the roadway surrounding rock is mainly thin-layered, and a large number of fissures are developed in the surrounding rock near the surface of the roadway. Through the in situ stress test, it is concluded that the in situ stress field of Jiahe Mine is dominated by horizontal stress, the minimum principal stress is vertical principal stress, and the lateral pressure coefficient exceeds 2, causing serious damage to the two sides of the roadway. Using the gas drainage drilling data of the 2254 east floor roadway and using SketchUp and Rhino software, the complex geological occurrence model of the coal seam above the roadway was constructed, and on this basis, the FLAC3D numerical calculation model was constructed. The analysis results show that the plastic zone of the roadway surrounding rock is large, the bolt anchoring section is within the plastic zone, and the control effect on the deformation of the roadway surrounding rock is poor. Therefore, a comprehensive support method of “reserved deformation, anchor rod, metal mesh, shotcrete, grouting anchor cable, and full-section grouting” is proposed. The monitoring results of the surrounding rock deformation of the roadway 12 months after construction show that the support scheme can effectively control the deformation of the roadway surrounding rock. The research results provide a reference for the stability control of similar roadways.


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## Body

## 1. Introduction

In recent years, my country’s national economy has developed rapidly, and energy demand has remained high. In 2021, my country’s total coal production will reach 4.13 billion tons, and the proportion of coal’s primary energy will remain stable at more than 50% for a long time. Coal mines in southern China generally have complex occurrence conditions and poor mining conditions. Therefore, there are many factors affecting roadway deformation during the mining process, and the deformation mechanism is unclear. At present, scholars at home and abroad have carried out certain research on roadway deformation and support under complex conditions. Chen et al. [1] believe that there is a difference in the original rock stress between the coal seam under complex geological conditions and the coal seam under ordinary conditions. Wang et al. [2] believed that the impact tendency of coal rock mass in complex geological occurrence environments is strong. Liang et al. [3] used three-dimensional microseismic monitoring technology to detect the microseismic activities inside the tunnel surrounding rock with complex geological conditions and high ground stress. Yanlin et al. [4] analyze the uncertainty, complexity, and diversity of geological bodies and enumerate the complex geological conditions that control the spatial form and physical distribution of geological bodies. Li et al. [5] constructed a multicoal seam logging interpretation model. Liang et al. [6] discussed the occurrence mechanism and key technologies for the prevention and control of coal-rock gas dynamic disasters in a magmatic rock occurrence environment. Jiang et al. [7] studied the failure law of thin coal seams under complex conditions. Wang et al. [8] evaluated the stability of goaf with complex landforms and multiple coal seams. Zheng [9] studied the maximum cutting height of complex thick coal seam conditions. Lu et al. [10] believed that the plastic failure of complex coal bodies will lead to a large amount of gas release. Wang et al. [11] studied the factors of back-arching instability of roadway floor under complex conditions. Yuan et al. [12] believed that the deformation symmetry of the roadway surrounding rock is poor under complex conditions, so asymmetric support is a better choice. Oparin et al. [13] introduced the comprehensive method and results of deformation wave monitoring of temporary lining of railway tunnels under complex geological conditions in southern Western Siberia. Wu et al. [14] discussed the influence of mud parameters and drilling speed on the quality of hole formation under complex geological conditions and obtained effective drilling speed and economical and reasonable mud parameters. To simulate the seismic performance of subway tunnels under complex geological conditions excited by oblique-incidence earthquakes, Gao et al. [15] established an overall numerical model of subway tunnels and foundations. Jianbing et al. [16] established a local elastic foundation model, studied the mechanical properties of the tunnel filling karst effect on the lining structure, and derived the calculation formulas for the displacement, bending moment, and shear force of the equivalent beam. Li [17] used the numerical simulation method to reveal the variation law of borehole stress after coal seam cutting through drilling and bedding drilling. Liaojun et al. [18] proposed a stability analysis method that can reflect the complex geological structure of the dam foundation. To realize safe and efficient mining in thin coal seams with complex geological conditions, Tongbin et al. [19] proposed an innovative pressure-regulated mining method. Yu et al. [20, 21] clarified the main points of support for weak semicoal rock roadways and proposed the idea of strengthening the lateral support of surrounding rock and improving the overall mechanical strength of such roadways. Wang et al. [22–24] found that the tunnel floor heave was inversely proportional to the strength of the two coal bodies and proposed that the deformation space should be reserved in advance, and the supporting structure should be adapted to it. Cui et al. [25] studied the deformation and stress distribution law of mining roadway and obtained the influence law of mining stress. The above studies have proposed the failure mechanism and stability control of the roadway surrounding rock under the condition of complex occurrence of coal gobs, but the roadway surrounding rock of the southern coal mine is weak, and the coal body occurrence conditions are complex. The existing roadway surrounding rock control measures still have certain limitations and cannot fully guide the stability control of the same type of roadway surrounding rock. Therefore, this paper takes the 2254 east floor roadway of Puxi Well as the object, analyzes the deformation mechanism of floor roadway under the condition of complex coal body, and proposes a comprehensive support method. This method is suitable for most of the same type of roadway and can effectively make up for the limitations of the existing roadway surrounding rock stability control measures.


## 2. Engineering Background

The Puxi mine of Jiahe Mining Company is located in the northeast of Jiahe County. The geological conditions of the rock formations are complex. The structures are mainly divided into three groups: north-south, north-west, and east-west. The overall mine rock formation is an undulating and sinuous syncline-anticline composite structure. The 2254 working face starts from the south wing of the 22 mining area -80 horizontal transport stone gate, north to the 22 mining area F63 fault boundary, east to the 11511 working face that has been mined, and west to the 22 mining area -105 horizontal gas drainage control area edge. Excavating along the V coal floor, the coal seam structure is complex and bumpy and can be exploited locally. The comprehensive histogram of this area is shown in Figure1. 2254 east floor roadway has a semicircular arch structure with a cross-section of 2600mm×2700mm and adopts the form of bolt mesh spraying support, and the length of the bolt is 1800 mm.Figure 1
Comprehensive columnar diagram of roadway surrounding rock.


## 3. Analysis of Deformation and Failure Characteristics of Roadway Surrounding Rock

### 3.1. Analysis of Surface Failure Characteristics of Roadway Surrounding Rock

Under the original support scheme of the 2254 east floor roadway, the roof bends and sinks seriously, the local shotcrete layer falls off, the layered surrounding rock is broken, and the separation layer between the surrounding rock layers is common, and the anchor bolt is suspended, as shown in Figure2.Figure 2
Bending deformation of roadway roof.The two sides of the roadway are asymmetrically deformed, the left side of the roadway is relatively stable, and the right side of the roadway shows serious intrusion, chipping, cracking, and other phenomena. As shown in Figure3, the “S”-shaped extrusion and bending deformation occurred in the surrounding rock, which led to the collapse of the right side roadway and the exposed bolts, which seriously affected the normal passage of the roadway.Figure 3
Rib deformation of roadway.


### 3.2. Analysis of the Development of Fissures in Road Surrounding Rock

A borehole peeping instrument is used to record the internal damage of the roadway roof and the surrounding rock of the side. The peeping results of the surrounding rock at the side of the roadway show that (Figure4(a)) there are obviously broken zones of surrounding rock within two ranges of 0.3 m-0.4 m and 0.8 m-1.1 m from the surface of the roadway side. Single cracks developed at 0.5 m and 1.3 m. There is a cavity at 1.6 m-1.7 m, the whole is continuous at 1.8 m-3.0 m, and cracks develop at 3.0 m-3.2 m.Figure 4
Peek view of borehole roadway surrounding rock.
(a)
Rib(b)
RoofWhen the tunnel roof was drilled and peeped, there was a broken rock zone at 1.6 m of the drilling position, so only the surrounding rock data within 1.6 m of the roof was obtained (Figure4(b)). The results show that there are a lot of cracks in the surrounding rock at 0.1 m~0.6 m. The vertical and horizontal cracks can be seen at 0.8 m~0.9 m, the integrity of the surrounding rock within 0.9 m~1.1 m is good, and there are cracks at 1.2 m, 1.4 m, and 1.5 m.


### 3.3. Mineral Composition Analysis of Roadway Surrounding Rock

To further understand the characteristics of the roadway surrounding rock, we used an X-ray diffractometer to analyze its composition. The analysis results show that the mineral composition of the roadway surrounding rock is mainly siderite, with a content of 51.34%, followed by quartz 15.82, mica (15.38%), and kaolinite (9.52%), in addition to a small amount of anatase and feldspar. Among them, the nature of siderite is relatively unstable, and it is easy to be weathered or hydrolyzed in contact with air. Due to its special structure, the mica will become very fine debris under the influence of pressure or weathering. Kaolinite with a larger content is less hard, often in the form of soil or lumps. If the environment is humid or water enters, it will easily become ooze or even disintegrate.


### 3.4. Ground Stress Test

To grasp the stress environment of the roadway surrounding rock, we carried out the in situ stress test on the coal mine. The test results are shown in Table1.Table 1
Horizontal and vertical stress data table.
Test hole numberHorizontal principal stress/MPaVertical principal stress/MPaLateral pressure coefficientZL-18.66.51.3ZL-29.85.41.8Mean9.26.01.5The test results show that the maximum horizontal stress in the mining area is 1.8 times the vertical stress, and the horizontal principal stress direction is perpendicular to the roadway axis.


## 3.1. Analysis of Surface Failure Characteristics of Roadway Surrounding Rock

Under the original support scheme of the 2254 east floor roadway, the roof bends and sinks seriously, the local shotcrete layer falls off, the layered surrounding rock is broken, and the separation layer between the surrounding rock layers is common, and the anchor bolt is suspended, as shown in Figure2.Figure 2
Bending deformation of roadway roof.The two sides of the roadway are asymmetrically deformed, the left side of the roadway is relatively stable, and the right side of the roadway shows serious intrusion, chipping, cracking, and other phenomena. As shown in Figure3, the “S”-shaped extrusion and bending deformation occurred in the surrounding rock, which led to the collapse of the right side roadway and the exposed bolts, which seriously affected the normal passage of the roadway.Figure 3
Rib deformation of roadway.


## 3.2. Analysis of the Development of Fissures in Road Surrounding Rock

A borehole peeping instrument is used to record the internal damage of the roadway roof and the surrounding rock of the side. The peeping results of the surrounding rock at the side of the roadway show that (Figure4(a)) there are obviously broken zones of surrounding rock within two ranges of 0.3 m-0.4 m and 0.8 m-1.1 m from the surface of the roadway side. Single cracks developed at 0.5 m and 1.3 m. There is a cavity at 1.6 m-1.7 m, the whole is continuous at 1.8 m-3.0 m, and cracks develop at 3.0 m-3.2 m.Figure 4
Peek view of borehole roadway surrounding rock.
(a)
Rib(b)
RoofWhen the tunnel roof was drilled and peeped, there was a broken rock zone at 1.6 m of the drilling position, so only the surrounding rock data within 1.6 m of the roof was obtained (Figure4(b)). The results show that there are a lot of cracks in the surrounding rock at 0.1 m~0.6 m. The vertical and horizontal cracks can be seen at 0.8 m~0.9 m, the integrity of the surrounding rock within 0.9 m~1.1 m is good, and there are cracks at 1.2 m, 1.4 m, and 1.5 m.


## 3.3. Mineral Composition Analysis of Roadway Surrounding Rock

To further understand the characteristics of the roadway surrounding rock, we used an X-ray diffractometer to analyze its composition. The analysis results show that the mineral composition of the roadway surrounding rock is mainly siderite, with a content of 51.34%, followed by quartz 15.82, mica (15.38%), and kaolinite (9.52%), in addition to a small amount of anatase and feldspar. Among them, the nature of siderite is relatively unstable, and it is easy to be weathered or hydrolyzed in contact with air. Due to its special structure, the mica will become very fine debris under the influence of pressure or weathering. Kaolinite with a larger content is less hard, often in the form of soil or lumps. If the environment is humid or water enters, it will easily become ooze or even disintegrate.


## 3.4. Ground Stress Test

To grasp the stress environment of the roadway surrounding rock, we carried out the in situ stress test on the coal mine. The test results are shown in Table1.Table 1
Horizontal and vertical stress data table.
Test hole numberHorizontal principal stress/MPaVertical principal stress/MPaLateral pressure coefficientZL-18.66.51.3ZL-29.85.41.8Mean9.26.01.5The test results show that the maximum horizontal stress in the mining area is 1.8 times the vertical stress, and the horizontal principal stress direction is perpendicular to the roadway axis.


## 4. Stability Simulation Analysis of Floor Roadway and Optimization of Support Parameters

### 4.1. Building the Geological Model

In order to grasp the occurrence of the coal seam above the 2254 east floor roadway, using the parameter map of the gas drainage drilling results in the roadway (Figure5(a)), and using SketchUp combined with Rhino software, the coal seam fitting lower surface and upper surface were established.Figure 5
(a) Gas drainage drilling results of 15 groups in the 2254 east floor roadway. (b) Each borehole rock mass and coal seam foundation model. (c) Coal seam fitting surface and geological model.
(a)(b)(c)Firstly, the SketchUp software is used to determine the drilling angle according to the azimuth and inclination in each group of extraction drilling holes. Then, use the black curve, the red dotted line, and the yellow curve to draw the length of the rock mass, coal seam, and overtop, respectively, at the drilling position, so as to construct the rock mass and coal seam foundation model of each drilling hole above the roadway, as shown in Figure5(b).The coal length curves made by different hole numbers of each borehole are connected to establish the coal seam profile above each group of borehole roadways. Import the basic model built by SketchUp into Rhino, determine the relative position and unit of the imported model, and fit all the lower and upper surface points to the lower surface and upper surface of the coal seam. In order to make the surface closer to the point cloud, try to let the fitted surface pass more feature points, and set the most suitable fitting parameters through continuous debugging, so as to obtain the final model, as shown in Figure5(c).


### 4.2. Numerical Simulation Model and Related Calculation Parameters

The geological model established by Rhino was imported into FLAC3D, and the situation of the 2254 east floor roadway affected by the occurrence of complex coal seams was analyzed. According to the geological structure obtained from field exploration, the model is divided into six layers from top to bottom: medium-grained sandstone, siltstone, coal, sandy mudstone, roadway, and argillaceous sandstone. The dimension of the straight wall semicircular arch roadway is2600mm×2700mm (width×height). The roadway was originally supported by anchors, nets, and spray. The anchor rod is Φ16×1800mm full-threaded steel, and the anchor rod is anchored by a resin coil. The model adopts the Cartesian coordinate system, and the generated mesh is divided into 670,020 elements and 114,705 nodes. At the same time, to ensure more accurate simulation, the surrounding rock within 6 m around the roadway is encrypted, as shown in Figure 6.Figure 6
Numerical simulation model.Model boundary conditions: lower boundary: displacement is zero; left, right, front, and rear boundaries: horizontal displacement is zero; upper boundary: free boundary, vertical stress 6 MPa; model lateral pressure coefficient is set to 2. The model adopts the Mohr-Coulomb yield criterion. The mechanical parameters of the surrounding rock are shown in Table2.Table 2
Numerical calculation parameters.
Surrounding rock typeBulk modulus/MPaShear modulus/MPaRock mass density/kg∙m3Internal friction angle/(°)Cohesive force/MPaMedium-grained sandstone21001.322750323.50Siltstone32002.702650302.30Coal2700.162235220.75Sandy mudstone9000.012600252.95


### 4.3. Mechanical Analysis of the Stability of Floor Roadway under the Influence of Upper Working Face Mining

To study the stability of 2254 east floor roadway under the influence of working face mining, we must first understand the distribution and transmission law of lateral support pressure on the working face floor. By referring to the literature [26], the maximum principal stress transfer law of the working face was obtained, and the contour map of the maximum principal stress distribution of the floor strata was obtained, as shown in Figure 7.Figure 7
Contour map of maximum principal stress distribution of floor rock stratum.Mining of the coal seam above the floor roadway leads to a regular transfer of the maximum principal stress of the rock formation. The main influencing factors include the vertical distance between the roadway and the upper working face and the horizontal distance between the coal wall edge of the roadway working face. Therefore, after the floor roadway is mined in complex coal seams, the stress distribution of the floor strata in the working face is closely related to the deployment position of the roadway. The difference in the spatial position between the goaf and the roadway makes the stress increase coefficient of the floor roadway vary greatly. According to the elastic mechanics’ analysis, the transfer law of the stress generated during the mining of the working face in the floor rock can be obtained.As shown in Figure8, the floor rock formation at the boundary of the gob is regarded as a semi-infinite plane body. Under the action of the concentrated load F along the positive direction of Z in the figure, the polar coordinate expression of the mining stress at any position Mρ,θ in the floor stratum is:
(1)σz=−2Fπcos3θρ,(2)σx=−2Fπsin2θcosθρ,(3)τxz=−2Fπsinθcos2θρ.Figure 8
Semi-infinite planar body of floor rock under the action of loadF.In the formula,σx, σz, and τxz represent the stress components in three directions, respectively, which are converted into a rectangular coordinate system and expressed as:
(4)σz=−2Fπx3x2+y22,(5)σx=−2Fπxy2x2+y22.Based on the above analysis, it can be concluded that the maximum vertical stress generated by mining is distributed on the lower axis of the load center, which is negatively related to the depth of the rock mass, that is, the smaller the depth, the greater the stress. At the same time, the closer to the axis, the greater the stress, and the easier it is for the surrounding rock to enter the plastic failure state.


### 4.4. Stability Simulation Analysis of 2254 East Floor Roadway Affected by Upper Working Face Mining

As shown in Figure9, after the working face is mined, the original stress balance state of the surrounding rock is broken, and the gravity of the overlying rock layer in the goaf is transferred to the front of the working face in the direction of the strike to form an advanced support pressure. In the direction of inclination, the pressure is transferred to both sides of the working face to form the support pressure on both sides of the working face. A certain stress-increasing area is formed in the bottom plate of the working face.Figure 9
Cloud map of horizontal stress and vertical stress distribution of roadway under the influence of mining.
(a)
Horizontal stress(b)
Vertical stressFigure9(a) shows that the horizontal stress distribution on both sides is different, the right side has a large deformation stress reduction area with a depth of about 2.4 m, and the affected area of the right side is about 1.5 m. Figure 9(b) shows that the right side of the roadway with the vertical stress floor forms a stress concentration area, and at the same time, a deformation stress reduction area appears on the roof on the right side of the roadway.Figure10 shows that after the mining effect, the plastic zone in the roadway surrounding rock is large, and the plastic zone is asymmetrically distributed. The width of the surrounding rock plastic zone on the right side of the roadway is 3.8 m, the depth of the bottom plastic zone is 3.2 m, the depth of the roof plastic zone is 2.1 m, and the minimum plastic zone on the left side is 1.8 m. The depth of the plastic zone exceeds the length of the bolt, the effect of bolt support is weakened, and the deformation of the roadway surrounding rock cannot be effectively controlled.Figure 10
Distribution of plastic zone of the roadway under the influence of mining.To sum up, after the upper working face of the roadway is mined, due to the increase of support pressure caused by mining, the original support scheme cannot control the stability of the floor roadway surrounding rock.


### 4.5. 2254 East Floor Roadway Improvement Support Scheme Design and Effect Simulation Analysis

Based on the above analysis, a comprehensive support method of “reserved deformation, anchor rod, metal mesh, shotcrete, grouting anchor cable, and full-section grouting” is proposed for the entire roadway section. Ensure the overall stability of the 2254 east floor roadway and prevent the large deformation of its surrounding rock from affecting the normal production of the mine.As shown in Figure11, the anchor rod is made of BHRB500 Ф22mm×2200mm left-handed rebar without longitudinal reinforcement. Each anchor uses 2 rolls of K2330 resin anchoring agent, the anchoring force is not less than 70 KN, the row spacing between the anchors is 700mm×700mm, the metal mesh is Ф6mm, the grid is 100×100mm, and the specification is 1000mm×800mm. There must be anchors where the metal mesh joins the stubble and is close to the rock surface, and the length of the stubble between the meshes is not less than 100 mm. The thickness of sprayed concrete is 100 mm, and the strength is C20. Using a steel plate tray, it is made into a square with a side length of 150 mm and a thickness of 8 mm.Figure 11
Anchor cable support section view.The anchor cable adopts a grouting anchor cable withФ21mm×6100mm, and the distance and discharge are 1200×1400mm. The end of the anchor is anchoring the resin, and the length of the anchor is 1500 mm. Each anchor cable uses 5 rolls of Z2850 resin rolls, and the anchoring power is not less than 100 kn. The anchor tray uses a compressed pad device, which is superimposed from two pads. The specifications are 350mm×350mm×10mm and 150mm×150mm×10mm square pads. The large pads are inside and the small pads are outside.Select the roadway section of the vertical height of the working surface to 6.4 m as the research object. From the perspective of the stress concentration of the roadway surrounding rock, the development characteristics of the plastic area, the bottom plate of the lane roads, and the two groups of fenced rock displacement comparison the comparison of several aspects, the support effect of the optimization scheme is analyzed.As shown in Figure12, after the optimization support scheme is adopted, the peak stress peak of the surrounding rock level X is reduced compared with the original scheme. The distribution of horizontal stress on the sides of the surrounding rock is still asymmetric. The right side of the lane road is still greater than the left. The distribution of vertical stress is still the deformation stress release area on the right side of the top plate. In general, the stress peak value has a certain decrease than the original plan, which improves the overall stability of the surrounding rock.Figure 12
Optimize the horizontal stress and vertical stress distribution cloud chart after the optimization of the roadway.
(a)
Horizontal stress(b)
Vertical stressIt can be seen from Figure13 that after the optimization scheme is adopted, the degree of asymmetric distribution of the plastic zone of the roadway surrounding rock is reduced. Among them, the width of the surrounding rock plastic zone on the right side of the roadway is reduced to 2.1 m, and the depth of the bottom plastic zone is reduced to 1.85 m. At the same time, in the improved support scheme, the length of the bolt is increased to 2200 mm, and the anchoring sections of the bolt and the cable are located in the elastic zone of the surrounding rock outside the plastic zone of the surrounding rock.Figure 13
Distribution of plastic zone of the roadway under optimization scheme.


## 4.1. Building the Geological Model

In order to grasp the occurrence of the coal seam above the 2254 east floor roadway, using the parameter map of the gas drainage drilling results in the roadway (Figure5(a)), and using SketchUp combined with Rhino software, the coal seam fitting lower surface and upper surface were established.Figure 5
(a) Gas drainage drilling results of 15 groups in the 2254 east floor roadway. (b) Each borehole rock mass and coal seam foundation model. (c) Coal seam fitting surface and geological model.
(a)(b)(c)Firstly, the SketchUp software is used to determine the drilling angle according to the azimuth and inclination in each group of extraction drilling holes. Then, use the black curve, the red dotted line, and the yellow curve to draw the length of the rock mass, coal seam, and overtop, respectively, at the drilling position, so as to construct the rock mass and coal seam foundation model of each drilling hole above the roadway, as shown in Figure5(b).The coal length curves made by different hole numbers of each borehole are connected to establish the coal seam profile above each group of borehole roadways. Import the basic model built by SketchUp into Rhino, determine the relative position and unit of the imported model, and fit all the lower and upper surface points to the lower surface and upper surface of the coal seam. In order to make the surface closer to the point cloud, try to let the fitted surface pass more feature points, and set the most suitable fitting parameters through continuous debugging, so as to obtain the final model, as shown in Figure5(c).


## 4.2. Numerical Simulation Model and Related Calculation Parameters

The geological model established by Rhino was imported into FLAC3D, and the situation of the 2254 east floor roadway affected by the occurrence of complex coal seams was analyzed. According to the geological structure obtained from field exploration, the model is divided into six layers from top to bottom: medium-grained sandstone, siltstone, coal, sandy mudstone, roadway, and argillaceous sandstone. The dimension of the straight wall semicircular arch roadway is2600mm×2700mm (width×height). The roadway was originally supported by anchors, nets, and spray. The anchor rod is Φ16×1800mm full-threaded steel, and the anchor rod is anchored by a resin coil. The model adopts the Cartesian coordinate system, and the generated mesh is divided into 670,020 elements and 114,705 nodes. At the same time, to ensure more accurate simulation, the surrounding rock within 6 m around the roadway is encrypted, as shown in Figure 6.Figure 6
Numerical simulation model.Model boundary conditions: lower boundary: displacement is zero; left, right, front, and rear boundaries: horizontal displacement is zero; upper boundary: free boundary, vertical stress 6 MPa; model lateral pressure coefficient is set to 2. The model adopts the Mohr-Coulomb yield criterion. The mechanical parameters of the surrounding rock are shown in Table2.Table 2
Numerical calculation parameters.
Surrounding rock typeBulk modulus/MPaShear modulus/MPaRock mass density/kg∙m3Internal friction angle/(°)Cohesive force/MPaMedium-grained sandstone21001.322750323.50Siltstone32002.702650302.30Coal2700.162235220.75Sandy mudstone9000.012600252.95


## 4.3. Mechanical Analysis of the Stability of Floor Roadway under the Influence of Upper Working Face Mining

To study the stability of 2254 east floor roadway under the influence of working face mining, we must first understand the distribution and transmission law of lateral support pressure on the working face floor. By referring to the literature [26], the maximum principal stress transfer law of the working face was obtained, and the contour map of the maximum principal stress distribution of the floor strata was obtained, as shown in Figure 7.Figure 7
Contour map of maximum principal stress distribution of floor rock stratum.Mining of the coal seam above the floor roadway leads to a regular transfer of the maximum principal stress of the rock formation. The main influencing factors include the vertical distance between the roadway and the upper working face and the horizontal distance between the coal wall edge of the roadway working face. Therefore, after the floor roadway is mined in complex coal seams, the stress distribution of the floor strata in the working face is closely related to the deployment position of the roadway. The difference in the spatial position between the goaf and the roadway makes the stress increase coefficient of the floor roadway vary greatly. According to the elastic mechanics’ analysis, the transfer law of the stress generated during the mining of the working face in the floor rock can be obtained.As shown in Figure8, the floor rock formation at the boundary of the gob is regarded as a semi-infinite plane body. Under the action of the concentrated load F along the positive direction of Z in the figure, the polar coordinate expression of the mining stress at any position Mρ,θ in the floor stratum is:
(1)σz=−2Fπcos3θρ,(2)σx=−2Fπsin2θcosθρ,(3)τxz=−2Fπsinθcos2θρ.Figure 8
Semi-infinite planar body of floor rock under the action of loadF.In the formula,σx, σz, and τxz represent the stress components in three directions, respectively, which are converted into a rectangular coordinate system and expressed as:
(4)σz=−2Fπx3x2+y22,(5)σx=−2Fπxy2x2+y22.Based on the above analysis, it can be concluded that the maximum vertical stress generated by mining is distributed on the lower axis of the load center, which is negatively related to the depth of the rock mass, that is, the smaller the depth, the greater the stress. At the same time, the closer to the axis, the greater the stress, and the easier it is for the surrounding rock to enter the plastic failure state.


## 4.4. Stability Simulation Analysis of 2254 East Floor Roadway Affected by Upper Working Face Mining

As shown in Figure9, after the working face is mined, the original stress balance state of the surrounding rock is broken, and the gravity of the overlying rock layer in the goaf is transferred to the front of the working face in the direction of the strike to form an advanced support pressure. In the direction of inclination, the pressure is transferred to both sides of the working face to form the support pressure on both sides of the working face. A certain stress-increasing area is formed in the bottom plate of the working face.Figure 9
Cloud map of horizontal stress and vertical stress distribution of roadway under the influence of mining.
(a)
Horizontal stress(b)
Vertical stressFigure9(a) shows that the horizontal stress distribution on both sides is different, the right side has a large deformation stress reduction area with a depth of about 2.4 m, and the affected area of the right side is about 1.5 m. Figure 9(b) shows that the right side of the roadway with the vertical stress floor forms a stress concentration area, and at the same time, a deformation stress reduction area appears on the roof on the right side of the roadway.Figure10 shows that after the mining effect, the plastic zone in the roadway surrounding rock is large, and the plastic zone is asymmetrically distributed. The width of the surrounding rock plastic zone on the right side of the roadway is 3.8 m, the depth of the bottom plastic zone is 3.2 m, the depth of the roof plastic zone is 2.1 m, and the minimum plastic zone on the left side is 1.8 m. The depth of the plastic zone exceeds the length of the bolt, the effect of bolt support is weakened, and the deformation of the roadway surrounding rock cannot be effectively controlled.Figure 10
Distribution of plastic zone of the roadway under the influence of mining.To sum up, after the upper working face of the roadway is mined, due to the increase of support pressure caused by mining, the original support scheme cannot control the stability of the floor roadway surrounding rock.


## 4.5. 2254 East Floor Roadway Improvement Support Scheme Design and Effect Simulation Analysis

Based on the above analysis, a comprehensive support method of “reserved deformation, anchor rod, metal mesh, shotcrete, grouting anchor cable, and full-section grouting” is proposed for the entire roadway section. Ensure the overall stability of the 2254 east floor roadway and prevent the large deformation of its surrounding rock from affecting the normal production of the mine.As shown in Figure11, the anchor rod is made of BHRB500 Ф22mm×2200mm left-handed rebar without longitudinal reinforcement. Each anchor uses 2 rolls of K2330 resin anchoring agent, the anchoring force is not less than 70 KN, the row spacing between the anchors is 700mm×700mm, the metal mesh is Ф6mm, the grid is 100×100mm, and the specification is 1000mm×800mm. There must be anchors where the metal mesh joins the stubble and is close to the rock surface, and the length of the stubble between the meshes is not less than 100 mm. The thickness of sprayed concrete is 100 mm, and the strength is C20. Using a steel plate tray, it is made into a square with a side length of 150 mm and a thickness of 8 mm.Figure 11
Anchor cable support section view.The anchor cable adopts a grouting anchor cable withФ21mm×6100mm, and the distance and discharge are 1200×1400mm. The end of the anchor is anchoring the resin, and the length of the anchor is 1500 mm. Each anchor cable uses 5 rolls of Z2850 resin rolls, and the anchoring power is not less than 100 kn. The anchor tray uses a compressed pad device, which is superimposed from two pads. The specifications are 350mm×350mm×10mm and 150mm×150mm×10mm square pads. The large pads are inside and the small pads are outside.Select the roadway section of the vertical height of the working surface to 6.4 m as the research object. From the perspective of the stress concentration of the roadway surrounding rock, the development characteristics of the plastic area, the bottom plate of the lane roads, and the two groups of fenced rock displacement comparison the comparison of several aspects, the support effect of the optimization scheme is analyzed.As shown in Figure12, after the optimization support scheme is adopted, the peak stress peak of the surrounding rock level X is reduced compared with the original scheme. The distribution of horizontal stress on the sides of the surrounding rock is still asymmetric. The right side of the lane road is still greater than the left. The distribution of vertical stress is still the deformation stress release area on the right side of the top plate. In general, the stress peak value has a certain decrease than the original plan, which improves the overall stability of the surrounding rock.Figure 12
Optimize the horizontal stress and vertical stress distribution cloud chart after the optimization of the roadway.
(a)
Horizontal stress(b)
Vertical stressIt can be seen from Figure13 that after the optimization scheme is adopted, the degree of asymmetric distribution of the plastic zone of the roadway surrounding rock is reduced. Among them, the width of the surrounding rock plastic zone on the right side of the roadway is reduced to 2.1 m, and the depth of the bottom plastic zone is reduced to 1.85 m. At the same time, in the improved support scheme, the length of the bolt is increased to 2200 mm, and the anchoring sections of the bolt and the cable are located in the elastic zone of the surrounding rock outside the plastic zone of the surrounding rock.Figure 13
Distribution of plastic zone of the roadway under optimization scheme.


## 5. Analysis and Evaluation of Measured Data of Roadway Deformation in New and Old Schemes

The new scheme was adopted for construction in 2254 east floor roadway. The construction layout is shown in Figure14. Observation points for the surface displacement of the roadway surrounding rock are set up within and outside the construction scope for data observation.Figure 14
Construction layout of the new scheme.


### 5.1. Observation Effect Analysis

The deformation of the roadway surrounding rock (the height of the roof and the floor and the width of the two sides) has been observed for one year. Figure15(a) is the roadway diagram after the construction of the new scheme, and Figure 15(b) is the roadway deformation diagram without the implementation of the new scheme.Figure 15
Comparison picture of deformation of roadway.
(a)
New scheme(b)
Without the new schemeThe following is a comparison chart of the height changes of the roof and floor at different monitoring points of the old and new schemes.Figures16(a) and 16(b) show that under the impact of mining, the top board of different observations has a certain degree of sinking. Point 4 (the construction section of the new plan) has the smallest sinking volume, and the height of the top board drops from the original 2540 mm to 2495 mm, a decrease of 45 mm. Point 5 (old plan construction section): the height of the top board dropped from the original 2200 mm to 1850 mm, the largest sinking volume, reaching 350 mm. The height of the top board of point 3 (the construction section of the old plan) dropped from 2300 mm to 2000 mm, sinking 300 mm. Point 2 (old plan construction section): top plate height dropped from 2400 mm to 2160 mm, sinking 240 mm. The height of the roadway at points 3 and 5 can no longer meet the normal walking needs. Pedestrians need to bow their heads to prevent head-butting. In addition, the roof height of no. 1 point (construction section of the old scheme, which is also located in the bottom plate of the goaf) is relatively small, from the original 2550 to 2450 mm, and the subsidence height is only 100 mm. Therefore, the support parameters should be further optimized for the roadway roof surrounding rock, especially the observation points 3 and 5.Figure 16
(a) Line chart of roadway roof and floor height change. (b) Column chart of roadway roof subsidence at different monitoring points.
(a)(b)


### 5.2. Analysis of Observation Effect on Both Sides of the Roadway

The monitoring data (Figure17) show that, under the influence of mining stress, the width on both sides of different monitoring points decreases to a certain extent. Point 4 (the construction section of the new scheme) has the smallest reduction in the width on both sides, from the original 2900 mm to 2850 mm, a reduction of 50 mm. Point 5 (construction section of the old scheme) was reduced from the original 2600 mm to 2460 mm, a reduction of 140 mm. Point 2 (construction section of the old scheme) has been reduced from the original 2640 mm to 2500 mm, a reduction of 140 mm. The no. 1 point (the construction section of the old scheme, which is also located in the bottom plate of the goaf) has been reduced from the original 2400 mm to 2300 mm, a reduction of 100 mm. No. 3 (the construction section of the old scheme): the width of the two gangs has been reduced from the original 2650 mm to 2550 mm, a reduction of 100 mm. Observation data show that the height of the top and bottom plates and the width of both sides of the roadway are significantly reduced after the new scheme is adopted, indicating that the new scheme can effectively control the deformation of the roadway.Figure 17
(a) Line chart of the height change of the two sides of the roadway. (b) The bar chart of the height change of the two sides of the roadway.
(a)(b)


## 5.1. Observation Effect Analysis

The deformation of the roadway surrounding rock (the height of the roof and the floor and the width of the two sides) has been observed for one year. Figure15(a) is the roadway diagram after the construction of the new scheme, and Figure 15(b) is the roadway deformation diagram without the implementation of the new scheme.Figure 15
Comparison picture of deformation of roadway.
(a)
New scheme(b)
Without the new schemeThe following is a comparison chart of the height changes of the roof and floor at different monitoring points of the old and new schemes.Figures16(a) and 16(b) show that under the impact of mining, the top board of different observations has a certain degree of sinking. Point 4 (the construction section of the new plan) has the smallest sinking volume, and the height of the top board drops from the original 2540 mm to 2495 mm, a decrease of 45 mm. Point 5 (old plan construction section): the height of the top board dropped from the original 2200 mm to 1850 mm, the largest sinking volume, reaching 350 mm. The height of the top board of point 3 (the construction section of the old plan) dropped from 2300 mm to 2000 mm, sinking 300 mm. Point 2 (old plan construction section): top plate height dropped from 2400 mm to 2160 mm, sinking 240 mm. The height of the roadway at points 3 and 5 can no longer meet the normal walking needs. Pedestrians need to bow their heads to prevent head-butting. In addition, the roof height of no. 1 point (construction section of the old scheme, which is also located in the bottom plate of the goaf) is relatively small, from the original 2550 to 2450 mm, and the subsidence height is only 100 mm. Therefore, the support parameters should be further optimized for the roadway roof surrounding rock, especially the observation points 3 and 5.Figure 16
(a) Line chart of roadway roof and floor height change. (b) Column chart of roadway roof subsidence at different monitoring points.
(a)(b)


## 5.2. Analysis of Observation Effect on Both Sides of the Roadway

The monitoring data (Figure17) show that, under the influence of mining stress, the width on both sides of different monitoring points decreases to a certain extent. Point 4 (the construction section of the new scheme) has the smallest reduction in the width on both sides, from the original 2900 mm to 2850 mm, a reduction of 50 mm. Point 5 (construction section of the old scheme) was reduced from the original 2600 mm to 2460 mm, a reduction of 140 mm. Point 2 (construction section of the old scheme) has been reduced from the original 2640 mm to 2500 mm, a reduction of 140 mm. The no. 1 point (the construction section of the old scheme, which is also located in the bottom plate of the goaf) has been reduced from the original 2400 mm to 2300 mm, a reduction of 100 mm. No. 3 (the construction section of the old scheme): the width of the two gangs has been reduced from the original 2650 mm to 2550 mm, a reduction of 100 mm. Observation data show that the height of the top and bottom plates and the width of both sides of the roadway are significantly reduced after the new scheme is adopted, indicating that the new scheme can effectively control the deformation of the roadway.Figure 17
(a) Line chart of the height change of the two sides of the roadway. (b) The bar chart of the height change of the two sides of the roadway.
(a)(b)


## 6. Conclusion

(1)
In situ stress test results show that the roadway surrounding rock is dominated by horizontal stress, and the maximum horizontal principal stress is perpendicular to the axial direction of the roadway, resulting in stress concentration in the roadway surrounding rock. At the same time, the surrounding rock composition of the roadway includes 9.52% kaolinite, 4.75% chlorite, 51.34% siderite, 15.38% mica, and other minerals, resulting in the relatively weak surrounding rock. It was found that the surrounding rock was damaged in the “M” type. At the same time, the drilling peek showed that there were a large number of cracks in the surrounding rock surface of the roadway within a range of 1.7 m(2)
Aiming at the complex occurrence of coal seams, a model was built using SketchUp combined with Rhino software and then imported into Flac 3D for calculation. The calculation results show that under the original scheme, the plastic zone of the roadway surrounding rock has a large development range, and the bolt is in the plastic zone, which cannot effectively control the deformation of the roadway(3)
A comprehensive support method of “reserved deformation, anchor rod, metal mesh, shotcrete, grouting anchor cable, and full-section grouting” is proposed for the 2254 east floor roadway, and numerical simulation is carried out. The simulation results show that under the new scheme, the approaching amount of the two sides of the roadway, the subsidence amount of the roof, and the bottom slump are greatly reduced(4)
The monitoring comparison of the deformation data of the roadway surrounding rock after the old and new support schemes shows that the deformation of the roadway roof and floor heights (45 mm) under the new scheme is significantly smaller than the maximum deformation (350 mm) of the old scheme. Under the new scheme, the deformation of the width of the two sides of the roadway (50 mm) is significantly smaller than the maximum deformation (140 mm) of the old scheme. The new scheme can meet the needs of normal work, and the overall support effect is good



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*Source: 7132188-2022-07-18.xml*