# Study on the Bearing Mechanism and Stability of Surrounding Rock in Original Roadway Filling and Nonpillar Tunneling

Authors: Wenbao Shi; Jucai Chang; Yan Li; Dong Li; Ru Hu
Journal: Geofluids
 (2023)

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## Abstract

This study is focused on the problem that with the increase of coal seam mining depth, it is difficult to continuously replace mining due to complex roadway layout and unreasonable stope layout. By taking the mining geological conditions of the 62210 fully mechanized mining face of Xinzhuangzi Coal Mine in Huainan mining area, China, as the background, it explores the stress characteristics of the original roadway filling body as well as the stress distribution and deformation characteristics of roadway surrounding rock in original roadway filling and nonpillar tunneling (ORFNPT) through theoretical analysis and numerical simulation. The following findings are obtained. The required strength for the filling body is primarily determined by two factors, i.e., the span of the hanging roof that lies over the filling body and the width of the filling body. The span of the hanging roof is positively correlated with the required strength of the filling body. However, when the width of the filling body reaches a certain value, its further increase fails to change the required strength of the filling body. Compared with gob-side entry driving with small coal pillars, when the ORFNPT technology is applied to the lower-section roadway, the peak stress position in the solid coal on the lower side of the roadway is closer to the roadway sides, and the filling body is of a much higher stress than the small coal pillars. Besides, the roadway surrounding rock undergoes milder deformation. According to the on-site application and measurement data, the roof-to-floor convergence and side-to-side displacement amounts of the roadway are about 89 mm and 58 mm during tunneling of the 62310 working face, and the two amounts are about 910 and 1,290 mm during recovery of the 62310 working face, respectively. This tunneling method achieves an excellent roadway control effect.


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## Body

## 1. Introduction

With the rapid economic development in China, the depth of coal mining is increasing year by year, and mining areas in eastern China have successively stepped into the stage of deep mining [1, 2]. In deep mining, the roadway surrounding rock is affected by complex stress fields such as high ground stress and intensive mining, which causes difficulties to the maintenance of roadway, thus hindering continuous mining of the stope [3, 4]. For old mining areas, stope replacement is faced with more difficulties induced by the increase in mining depth, the complex layout of mine roadway, and the unreasonable layout of stope [5]. Moreover, the retention of various coal pillars during coal mining not only notably affects the coal recovery rate but also leads to a large waste of resources.At present, the commonly used section roadway protection methods can mainly be classified into two categories: gob-side entry driving with small coal pillars (GSEDSCP) [6, 7] and gob-side entry retaining (GSER) [8, 9]. The former often leads to stress concentration in the underlying coal seams, which may induce geological disasters (such as large deformation, rock burst, and coal and gas outburst) in the surrounding rock of gob-side roadway during underlying coal seam recovery [10]. The latter is influenced by tunneling once and by mining twice, and generally, the roadway is barely functioning when used for the second time as a result of fast and serious surrounding rock deformation [11]. In order to ensure the stability of roadway in deep coal recovery, scholars have attached importance to the research on the support strength and deformation stress characteristics of the filling body in GSER in recent years. Kong et al. [12] deduced a calculation formula for the width of the roadway-side support. Qi et al. [13] and Huang et al. [14] studied the bearing characteristics of underground filling body and the support strength of working face and analyzed the technical difficulties and key points in the construction of solid filling mining support. Khaldoun et al. [15] investigated the movement process and deformation characteristics of roadway roof strata in a deep mine, determined a design principle for the support resistance of the filling body, and established a mathematical model for the support resistance and compression of the filling body. Chen et al. [16, 17] classified roadway roofs into three typical types, namely, thick immediate roof, thin immediate roof, and no immediate roof. Besides, they obtained formulae to calculate the support resistance of gob-side roadway for the above three types by superimposing the continuous layer model and considering the bearing effect of the coal on the retained roadway sides and the factors inducing roof collapse. Cheng et al. [18] analyzed the influence of the width of filling body and material characteristics on the stability of surrounding rock in fully mechanized caving roadways based on the key technology of surrounding rock control in such roadways and put forward measures to improve the deformation resistance capacity of the filling body. Feng and Zhang and Guo et al. [19, 20] simplified the deep coal seam roof as a rectangular “superimposed layer” without interlayer bonding force, analyzed the law of roadway roof activity in different periods, established a mechanical model of the relationship between roadway “surrounding rock and support” with the aid of elastic-plastic mechanical theory, and obtained a formula to calculate the support resistance. Kong et al. [21] analyzed the factors influencing roadway surrounding rock deformation in fully mechanized caving face, conducted in-depth analysis on the interaction mechanism between surrounding rock and filling body, and obtained a formula to calculate the support resistance of the filling body, which provided a theoretical basis for the on-site application of working face roadway. To solve the problem of gas accumulation in the upper corner of high-gas mines, Ma et al. [22, 23] proposed the secondary roadway technology, established a mechanical model of the key blocks at the ends, and obtained theoretical formulae to calculate the support resistance and the width of the filling body. Meng et al. and Mo et al. [24, 25] put forward a roadway technology of “concrete filling on the roadway sides + anchor-net-cable combined support in the roadway + anchor-cable reinforcement on the roadway sides” based on the activity law and deformation characteristics of roadway roof strata in a deep mine. Furthermore, they analyzed the relationship between roof separation and roof deformation and obtained the critical value of roof separation. Zhu et al. [26] established a mechanical model of the key block and the immediate roof and analyzed the mechanism of interaction between the key block and the roadway surrounding rock. The results revealed that the width of the filling body had a great influence on the stability of the gob-side roadway. Sun et al. [27] studied the deformation characteristics of fully mechanized caving roadways through similar simulation tests, concluding that the filling body should be of both a certain strength and a certain deformation resistance capacity. Seryakov [28] analyzed the deformation and stress distribution characteristics of roadway surrounding rock and proposed a design principle for deep roadway filling support by analyzing on-site mine pressure monitoring data. In summary, previous studies were focused on the bearing characteristics of the filling body of GSER as well as the interaction between the filling body and the surrounding rock. Constructing the filling body can effectively improve the stress state of the surrounding rock in the entire roadway and provide more support for the surrounding rock in the roadway roof. The research results serve as important reference for the application of the new method for recovery roadway tunneling, i.e., the original roadway filling and nonpillar tunneling (ORFNPT) technology.In this study, a mechanical model for original roadway filling under the influence of roof structure was established based on the basic law of roof stratum fracture and the key stratum theory. Next, the stress distribution and deformation characteristics of surrounding rock of ORFNPT were analyzed theoretically and numerically. Finally, on-site application was conducted in the 62210 working face of Xinzhuangzi Coal Mine.


## 2. Geological Overview

Xinzhuangzi Coal Mine is located in the west of Huainan City, China, and at the east foot of Bagong Mountain. The mine field stretches from the north side of Huaihe River to the south side of it. The 62210 fully mechanized working face of the mine, whose elevation ranges from -660 m to -770 m, has an average strike length of 1,285 m, an average dip length of 194 m, and an area of 249,290 m2. The average coal seam thickness is 1.0 m, and the coal seam dip angle lies in the range of 21°-30°, with an average of 25°. The corresponding overlying B11 and B8 coal seams have not been mined yet. The immediate roof belongs to sandy mudstone with a thickness of 2-12 m, and it is a fragile dark-gray thin layer. The main roof is fine sandstone with a thickness of 0-6 m. The immediate floor is fine sandstone with a thickness of 1-3.5 m, and it is a fragile gray thick layer. The main floor belongs to mudstone with a thickness of 1-4 m. The schematic diagram of geological overview of the working face is given in Figure 1.Figure 1
Schematic diagram of geological overview of the 62210 working face.


## 3. Force Analysis on the Filling Body in ORFNPT

### 3.1. Force Calculation on the Filling Body in Original Roadway Filling

The support resistance of the filling body in the gob-side roadway is usually calculated with the superimposed continuous layer model which considers the separation and dislocation between layers. In the model, each layer can be regarded as an independent continuous layer structure, and the layers are connected by distributed loads, so this model is faithful to the actual situation of stratum occurrence [29]. In addition, a mechanical structure model of the filling body support and roof was established with reference to the basic law of roof stratum fracture and the key stratum theory [30]. The support resistance of the filling body in the original roadway was studied through the block mechanical balance method. Figure 2(a) is the roof structure model with four sides supported, and Figure 2(b) is the one with three sides supported and one side free. During the first weighting, the roof in the fully mechanized mining face belongs to the model with four sides supported. q is the uniformly distributed load of roof; the load on ABCD surface after division by strip only concentrates on sections AB and CD, and the support resistance is Q.Figure 2
Calculation model of support strength of the filling body in the roadway.
(a)
Roof mechanical model for first weighting(b)
Roof radial strip mechanical modelThe calculation of the required strength of the filling body starts from the last stratum of the overlying strata. According to Figure2(a), the required strength of the filling body under a single overlying stratum is calculated as follows:
(1)SectionAB:∑Fy=0,aq1cosη+FA1−Q1a−FO1=0,∴FO1=q1a+L1−Q1a,where a is the width of the filling body (m), b is the span of immediate hanging roof above the filling body (m), η is the dip angle of coal seam, FA1 is the shear force generated downward by the fracture block at the rock fracture, and FA1=q1L1.During periodic weighting of the working face roof, the roof belongs to the roof structure model with three sides supported and one side free. For the above two models, a slat with unit width and a large load in the gob-side roadway is taken as the calculation unit (Figures3(a) and 3(b)).
(2)SectionOA:∑M=0,Mp1−M01+12q1cosηa+b2+q1cosηL1a+b−12Q1a2=0,where q1=γ1⋅h1cosη. Then, q1 is substituted in the expression of Q1:
(3)Q1=2Mp1−2M01+a+b2L1γ1h1cosη+a+bL12γ1h1cosηa2,where Mp1 is the ultimate bending moment of strata (kN·m), M01 is the bending moment of strata (kN·m), FA1 is the downward shear force generated by the rock fracture block of point A (kN), Q1 is the roof cutting strength (MPa), q1 is the rock mass concentration (MPa), γ1 is the bulk density of the first stratum above the filling body, h1 is the depth of the first stratum above the filling body (m), and L1 is the fracture characteristic size of the first stratum (m).Figure 3
Calculation model of support strength of the filling body during periodic weighting.
(a)
Mechanical model of roof during periodic weighting(b)
Schematic diagram of the mechanical model of roofUnder the second stratum fracture, the required strength of the filling objects isQ2, and the mechanical analysis on the stress state of the filling body is as follows:
(4)∑M=M01+M02+12Q2a2=Mp2+12γ2h2a+b+h1tanθ12cosη+12γ1h1a+b2cosη+FA2a+b+h1tanθ1+FA1a+b12Q2a2=12γ2h2a+b+h1tanθ12+12γ1h1a+b2+γ1h1L1a+b+γ2h2L2a+b+h1tanθ1cosη+Mp2−M01−M02=12∑i=12γ1h1a+b+∑j=0i=1hjtanθj2cosη+∑i=12FAia+b+∑j=0i=1hjtanθj+Mp2−∑i=12M0i.The strengths of the filling body under one layer or two layers of overlying strata are obtained through mechanical analysis on the first and second layers of the overlying strata of the filling body. Similarly, the required strengthQm of the filling body under the fracture of the m-th stratum above the filling body can be obtained as follows:Them-th layer of roof:
(5)Qm=∑i=1mγihia+b+∑j=0i=1hjtanθj2cosηa2+2Mpma2−2∑i=1mM0ia2+2∑i=1mFAia+b+∑j=0i=1hjtanθja2=∑i=1mγihi1+ba+1a∑j=0i=1hjtanθj2cosη+∑i=1mFAi1a+ba2+1a2∑j=0i=1hjtanθj+2Mpma2−2∑i=1mM0ia2,where Mpm is the ultimate bending moment under the fracture in the m-th stratum (kN·m), θj is the fracture angle of roof strata, and FAi is the shear force generated downward by fracture block of overlying strata (kN).It can be seen from Equation (5) that under certain geological conditions of the working face, the required strength (Q) of the filling body is inversely proportional to the width of the filling body (a), and the larger the width of the filling body is, the lower the required strength is. In addition, the required strength of the filling body is proportional to the span of the hanging roof (b) that lies over the filling body. A larger span of the roof needs a higher required strength of the filling body. When the surrounding rock and the supporting structure jointly form a stable support bearing system, the filling body is able to support the load of the overlying strata, thereby reducing the load on the filling body.


### 3.2. Engineering Example Calculation

With the geological conditions of the 62210 working face in Xinzhuangzi Coal Mine of Huainan mining area taken as the background, the stress characteristics of the original roadway filling body were calculated under the widths of 1 m, 2 m, 3 m, 5 m, and 7 m, respectively.According to on-site investigation, the immediate roof of the 62210 working face is sandy mudstone with a thickness of 6 m. The specific parameters are as follows: the bulk density (γ) is 25,100, and the fracture angle of strata is about 25° (measured on site). Under the limit conditions, the ultimate bending moment of strata is considered equal to the bending moment of the strata, i.e., M01=MP1. The working resistance of the filling body can be calculated by Equation (4) under different widths of the filling body. Figure 4 shows the relationship between the width of the filling body and the required strength of it.Figure 4
Stress variation curves of the filling body at different widths.Figure4 reveals the following phenomenon: as the width of the filling body increases from 1 m to 2 m, the required strength plunges dramatically at a high rate. From 2 m to 3 m, the required strength continues to drop at a high rate. From 3 m to 7 m, the required strength corresponds to a low decline rate. The width of the filling body plays an important role in the required strength of them. Considering the technical and economic benefits, the reasonable width of original roadway filling in the 62210 working face is 3 m. Therefore, the filling materials and filling process can be selected according to the required strength.


## 3.1. Force Calculation on the Filling Body in Original Roadway Filling

The support resistance of the filling body in the gob-side roadway is usually calculated with the superimposed continuous layer model which considers the separation and dislocation between layers. In the model, each layer can be regarded as an independent continuous layer structure, and the layers are connected by distributed loads, so this model is faithful to the actual situation of stratum occurrence [29]. In addition, a mechanical structure model of the filling body support and roof was established with reference to the basic law of roof stratum fracture and the key stratum theory [30]. The support resistance of the filling body in the original roadway was studied through the block mechanical balance method. Figure 2(a) is the roof structure model with four sides supported, and Figure 2(b) is the one with three sides supported and one side free. During the first weighting, the roof in the fully mechanized mining face belongs to the model with four sides supported. q is the uniformly distributed load of roof; the load on ABCD surface after division by strip only concentrates on sections AB and CD, and the support resistance is Q.Figure 2
Calculation model of support strength of the filling body in the roadway.
(a)
Roof mechanical model for first weighting(b)
Roof radial strip mechanical modelThe calculation of the required strength of the filling body starts from the last stratum of the overlying strata. According to Figure2(a), the required strength of the filling body under a single overlying stratum is calculated as follows:
(1)SectionAB:∑Fy=0,aq1cosη+FA1−Q1a−FO1=0,∴FO1=q1a+L1−Q1a,where a is the width of the filling body (m), b is the span of immediate hanging roof above the filling body (m), η is the dip angle of coal seam, FA1 is the shear force generated downward by the fracture block at the rock fracture, and FA1=q1L1.During periodic weighting of the working face roof, the roof belongs to the roof structure model with three sides supported and one side free. For the above two models, a slat with unit width and a large load in the gob-side roadway is taken as the calculation unit (Figures3(a) and 3(b)).
(2)SectionOA:∑M=0,Mp1−M01+12q1cosηa+b2+q1cosηL1a+b−12Q1a2=0,where q1=γ1⋅h1cosη. Then, q1 is substituted in the expression of Q1:
(3)Q1=2Mp1−2M01+a+b2L1γ1h1cosη+a+bL12γ1h1cosηa2,where Mp1 is the ultimate bending moment of strata (kN·m), M01 is the bending moment of strata (kN·m), FA1 is the downward shear force generated by the rock fracture block of point A (kN), Q1 is the roof cutting strength (MPa), q1 is the rock mass concentration (MPa), γ1 is the bulk density of the first stratum above the filling body, h1 is the depth of the first stratum above the filling body (m), and L1 is the fracture characteristic size of the first stratum (m).Figure 3
Calculation model of support strength of the filling body during periodic weighting.
(a)
Mechanical model of roof during periodic weighting(b)
Schematic diagram of the mechanical model of roofUnder the second stratum fracture, the required strength of the filling objects isQ2, and the mechanical analysis on the stress state of the filling body is as follows:
(4)∑M=M01+M02+12Q2a2=Mp2+12γ2h2a+b+h1tanθ12cosη+12γ1h1a+b2cosη+FA2a+b+h1tanθ1+FA1a+b12Q2a2=12γ2h2a+b+h1tanθ12+12γ1h1a+b2+γ1h1L1a+b+γ2h2L2a+b+h1tanθ1cosη+Mp2−M01−M02=12∑i=12γ1h1a+b+∑j=0i=1hjtanθj2cosη+∑i=12FAia+b+∑j=0i=1hjtanθj+Mp2−∑i=12M0i.The strengths of the filling body under one layer or two layers of overlying strata are obtained through mechanical analysis on the first and second layers of the overlying strata of the filling body. Similarly, the required strengthQm of the filling body under the fracture of the m-th stratum above the filling body can be obtained as follows:Them-th layer of roof:
(5)Qm=∑i=1mγihia+b+∑j=0i=1hjtanθj2cosηa2+2Mpma2−2∑i=1mM0ia2+2∑i=1mFAia+b+∑j=0i=1hjtanθja2=∑i=1mγihi1+ba+1a∑j=0i=1hjtanθj2cosη+∑i=1mFAi1a+ba2+1a2∑j=0i=1hjtanθj+2Mpma2−2∑i=1mM0ia2,where Mpm is the ultimate bending moment under the fracture in the m-th stratum (kN·m), θj is the fracture angle of roof strata, and FAi is the shear force generated downward by fracture block of overlying strata (kN).It can be seen from Equation (5) that under certain geological conditions of the working face, the required strength (Q) of the filling body is inversely proportional to the width of the filling body (a), and the larger the width of the filling body is, the lower the required strength is. In addition, the required strength of the filling body is proportional to the span of the hanging roof (b) that lies over the filling body. A larger span of the roof needs a higher required strength of the filling body. When the surrounding rock and the supporting structure jointly form a stable support bearing system, the filling body is able to support the load of the overlying strata, thereby reducing the load on the filling body.


## 3.2. Engineering Example Calculation

With the geological conditions of the 62210 working face in Xinzhuangzi Coal Mine of Huainan mining area taken as the background, the stress characteristics of the original roadway filling body were calculated under the widths of 1 m, 2 m, 3 m, 5 m, and 7 m, respectively.According to on-site investigation, the immediate roof of the 62210 working face is sandy mudstone with a thickness of 6 m. The specific parameters are as follows: the bulk density (γ) is 25,100, and the fracture angle of strata is about 25° (measured on site). Under the limit conditions, the ultimate bending moment of strata is considered equal to the bending moment of the strata, i.e., M01=MP1. The working resistance of the filling body can be calculated by Equation (4) under different widths of the filling body. Figure 4 shows the relationship between the width of the filling body and the required strength of it.Figure 4
Stress variation curves of the filling body at different widths.Figure4 reveals the following phenomenon: as the width of the filling body increases from 1 m to 2 m, the required strength plunges dramatically at a high rate. From 2 m to 3 m, the required strength continues to drop at a high rate. From 3 m to 7 m, the required strength corresponds to a low decline rate. The width of the filling body plays an important role in the required strength of them. Considering the technical and economic benefits, the reasonable width of original roadway filling in the 62210 working face is 3 m. Therefore, the filling materials and filling process can be selected according to the required strength.


## 4. Study on the Stability of Recovery Roadways

### 4.1. Tunneling Method of the B10 Coal Recovery Roadway

The B10 coal seam, which serves as the key protective seam of Xinzhuangzi Coal Mine, is mined for the purpose of relieving the pressure and protecting the overlying B11 coal seam and the underlying B8 coal seam. Therefore, in the mining process, the retention of large coal pillars for roadway protection should be minimized to avoid the influence of stress concentration on the underlying coal seam. Conventionally, the roadway layout often adopts GSEDSCP or GSER. GSEDSCP reduces the width of coal pillar, but it still induces stress concentration to the underlying coal seam. GSER does not involve the retention of coal pillars, but the retained roadway is barely functioning when used for the second time because it has suffered the disturbance of tunneling once and mining twice. On-site practice has also proved the poor stability of surrounding rock and a high maintenance cost when the GSER method is applied to roadway layout. Hence, according to the mine production and geological conditions, the ORFNPT technology is proposed here to arrange roadways in the stope.


### 4.2. Numerical Analysis

To analyze the stability characteristics of surrounding rock in ORFNPT and solve practical problems, a calculation model was established through large-scale finite difference software FLAC3D by taking the 62210 working face as the background. Some parameters of the model are listed as follows: strike length 320 m, dip length 230 m, height 248 m, and average coal seam thickness 1 m. The four sides and bottom of the model were fixed. Besides, a 13.8 MPa vertical stress was applied to the top of the model to simulate the mass of the overlying strata, and the lateral pressure coefficient was 1.0. The mechanical parameters of coal rock are given in Table1.Table 1
Mechanical parameters of coal rock involved in the simulation.
RockRock thickness (m)Elastic modulusE (GPa)Poisson’s ratioμCohesionC (MPa)Internal friction angleφ (o)Tensile strengthσT (MPa)Density d (kg/m3)Mudstone1.6622.50.223.21391.522660B11a coal1.075.130.331.15330.141387Sandy mudstone2.75.450.1472.16361.32520Sandstone2.413.70.132.02401.032497Carbon mudstone0.288.850.261.2300.6052461B11 coal0.754.910.341.35330.1321355Sandstone1.313.70.132.03401.032497Mudstone5.9516.80.222.22331.512484Siltstone5.741.810.2322.51371.222465Sandy mudstone6.125.450.152.26361.232533B10 coal15.50.241.52331.231465Mudstone5.116.80.222.22331.512484Sandy mudstone7.675.450.1472.16361.32520The four sides of the model were only a horizontal displacement boundary, and the floor was a fixed boundary. The simulated buried depth of the coal seam ranged from -660 m to -770 m, and a uniformly distributed load was applied to the top boundary of the model. The constitutive relation of roadway surrounding rock followed the modified Mohr-Coulomb criterion. The numerical simulation included two parts: (1) the vertical stress distribution characteristics of roadway surrounding rock in ORFNPT and GSEDSCP and (2) the deformation and displacement characteristics of roadway surrounding rock in ORFNPT and GSEDSCP.The simulation was conducted using single-factor analysis. For ORFNPT, the width and strength of the filling body were 3 m and 6.6 MPa, respectively. For GSER, the width of the coal pillar was 7 m.


#### 4.2.1. Stress Distribution Law of Roadway Surrounding Rock in Lower-Section Tunneling

The stress state of roadway surrounding rock is an important indicator to evaluate the stability of roadway, yet the stress characteristics of the coal rock for roadway protection in the gob-side roadway are more crucial for the success of the retained gob-side roadway. Therefore, numerical simulation was conducted on the above two tunneling methods. The yielded stress distribution nephograms is given in Figures5 and 6, and the stress variation curves are displayed in Figure 7. It can be known that under the same geological conditions, for different tunneling methods, the distribution characteristics of the roadway surrounding rock of the excavated roadway differ notably. For ORFNPT, the peak stress of the solid coal side is located at the position of about 3 m in the coal; the peak stress is about 41.5 MPa; the stress concentration coefficient is about 1.74, and the stress on the filling body is about 21 MPa. For GSEDSCP, the stress peak in the solid coal on the lower side of the roadway is about 4.25 m in the coal; the stress peak is around 43.2 MPa; and the stress concentration coefficient is 1.82. Besides, a stress concentration core (range about 1 m and stress peak about 17.2 MPa) exists in the small coal pillar. Compared with GSEDSCP, ORFNPT can protect the roadway tunneling in the lower section more effectively. Its peak stress of the solid coal on the lower side of the roadway is closer to the roadway sides, and the stress of the filling body is much higher than that of the small coal pillar, indicating that the filling body can support the overlying roof better and can bear more load from the overlying strata.Figure 5
Nephogram of vertical stress distribution of surrounding rock in ORFNPT along the coal seam dip.Figure 6
Nephogram of vertical stress distribution of surrounding rock in GSEDSCP along the coal seam dip.Figure 7
Stress distribution curves of roadway surrounding rock along the coal seam dip.


#### 4.2.2. Deformation and Displacement Distribution of Roadway Surrounding Rock in Lower-Section Tunneling

The displacement variations of roadway surrounding rock of the two tunneling methods were monitored (Figure8). For ORFNPT, the maximum displacement of the roof is 0.63 m; the maximum heave of the floor is about 0.35 m; and the horizontal displacement on the solid coal side is about 0.51 m. For GSEDSCP, the maximum displacement of the roof is about 1.21 m; the maximum heave of the floor is about 0.48 m; and the maximum horizontal displacement on the solid coal side is about 1.22 m.Figure 8
Distribution characteristics of roadway surrounding rock displacement in ORFNPT and GSEDSCP along the coal seam dip.
(a)
Variation curves of vertical displacement(b)
Variation curves of horizontal displacementThe comparison in Figure7 suggests that the stress field and displacement field of roadway surrounding rock differ under different strength of the filling body. Under a high strength of the filling body, the stress is more likely to be concentrated on the filling body and less likely to be concentrated on the roadway surrounding rock. Accordingly, the roof-to-floor convergence will be smaller, and the damage to the surrounding rock will be slighter. These data indicate that ORFNPT is of a milder roadway surrounding rock deformation than GSEDSCP and is thus more conducive to the maintenance and stability of the lower-section roadway.


## 4.1. Tunneling Method of the B10 Coal Recovery Roadway

The B10 coal seam, which serves as the key protective seam of Xinzhuangzi Coal Mine, is mined for the purpose of relieving the pressure and protecting the overlying B11 coal seam and the underlying B8 coal seam. Therefore, in the mining process, the retention of large coal pillars for roadway protection should be minimized to avoid the influence of stress concentration on the underlying coal seam. Conventionally, the roadway layout often adopts GSEDSCP or GSER. GSEDSCP reduces the width of coal pillar, but it still induces stress concentration to the underlying coal seam. GSER does not involve the retention of coal pillars, but the retained roadway is barely functioning when used for the second time because it has suffered the disturbance of tunneling once and mining twice. On-site practice has also proved the poor stability of surrounding rock and a high maintenance cost when the GSER method is applied to roadway layout. Hence, according to the mine production and geological conditions, the ORFNPT technology is proposed here to arrange roadways in the stope.


## 4.2. Numerical Analysis

To analyze the stability characteristics of surrounding rock in ORFNPT and solve practical problems, a calculation model was established through large-scale finite difference software FLAC3D by taking the 62210 working face as the background. Some parameters of the model are listed as follows: strike length 320 m, dip length 230 m, height 248 m, and average coal seam thickness 1 m. The four sides and bottom of the model were fixed. Besides, a 13.8 MPa vertical stress was applied to the top of the model to simulate the mass of the overlying strata, and the lateral pressure coefficient was 1.0. The mechanical parameters of coal rock are given in Table1.Table 1
Mechanical parameters of coal rock involved in the simulation.
RockRock thickness (m)Elastic modulusE (GPa)Poisson’s ratioμCohesionC (MPa)Internal friction angleφ (o)Tensile strengthσT (MPa)Density d (kg/m3)Mudstone1.6622.50.223.21391.522660B11a coal1.075.130.331.15330.141387Sandy mudstone2.75.450.1472.16361.32520Sandstone2.413.70.132.02401.032497Carbon mudstone0.288.850.261.2300.6052461B11 coal0.754.910.341.35330.1321355Sandstone1.313.70.132.03401.032497Mudstone5.9516.80.222.22331.512484Siltstone5.741.810.2322.51371.222465Sandy mudstone6.125.450.152.26361.232533B10 coal15.50.241.52331.231465Mudstone5.116.80.222.22331.512484Sandy mudstone7.675.450.1472.16361.32520The four sides of the model were only a horizontal displacement boundary, and the floor was a fixed boundary. The simulated buried depth of the coal seam ranged from -660 m to -770 m, and a uniformly distributed load was applied to the top boundary of the model. The constitutive relation of roadway surrounding rock followed the modified Mohr-Coulomb criterion. The numerical simulation included two parts: (1) the vertical stress distribution characteristics of roadway surrounding rock in ORFNPT and GSEDSCP and (2) the deformation and displacement characteristics of roadway surrounding rock in ORFNPT and GSEDSCP.The simulation was conducted using single-factor analysis. For ORFNPT, the width and strength of the filling body were 3 m and 6.6 MPa, respectively. For GSER, the width of the coal pillar was 7 m.


### 4.2.1. Stress Distribution Law of Roadway Surrounding Rock in Lower-Section Tunneling

The stress state of roadway surrounding rock is an important indicator to evaluate the stability of roadway, yet the stress characteristics of the coal rock for roadway protection in the gob-side roadway are more crucial for the success of the retained gob-side roadway. Therefore, numerical simulation was conducted on the above two tunneling methods. The yielded stress distribution nephograms is given in Figures5 and 6, and the stress variation curves are displayed in Figure 7. It can be known that under the same geological conditions, for different tunneling methods, the distribution characteristics of the roadway surrounding rock of the excavated roadway differ notably. For ORFNPT, the peak stress of the solid coal side is located at the position of about 3 m in the coal; the peak stress is about 41.5 MPa; the stress concentration coefficient is about 1.74, and the stress on the filling body is about 21 MPa. For GSEDSCP, the stress peak in the solid coal on the lower side of the roadway is about 4.25 m in the coal; the stress peak is around 43.2 MPa; and the stress concentration coefficient is 1.82. Besides, a stress concentration core (range about 1 m and stress peak about 17.2 MPa) exists in the small coal pillar. Compared with GSEDSCP, ORFNPT can protect the roadway tunneling in the lower section more effectively. Its peak stress of the solid coal on the lower side of the roadway is closer to the roadway sides, and the stress of the filling body is much higher than that of the small coal pillar, indicating that the filling body can support the overlying roof better and can bear more load from the overlying strata.Figure 5
Nephogram of vertical stress distribution of surrounding rock in ORFNPT along the coal seam dip.Figure 6
Nephogram of vertical stress distribution of surrounding rock in GSEDSCP along the coal seam dip.Figure 7
Stress distribution curves of roadway surrounding rock along the coal seam dip.


### 4.2.2. Deformation and Displacement Distribution of Roadway Surrounding Rock in Lower-Section Tunneling

The displacement variations of roadway surrounding rock of the two tunneling methods were monitored (Figure8). For ORFNPT, the maximum displacement of the roof is 0.63 m; the maximum heave of the floor is about 0.35 m; and the horizontal displacement on the solid coal side is about 0.51 m. For GSEDSCP, the maximum displacement of the roof is about 1.21 m; the maximum heave of the floor is about 0.48 m; and the maximum horizontal displacement on the solid coal side is about 1.22 m.Figure 8
Distribution characteristics of roadway surrounding rock displacement in ORFNPT and GSEDSCP along the coal seam dip.
(a)
Variation curves of vertical displacement(b)
Variation curves of horizontal displacementThe comparison in Figure7 suggests that the stress field and displacement field of roadway surrounding rock differ under different strength of the filling body. Under a high strength of the filling body, the stress is more likely to be concentrated on the filling body and less likely to be concentrated on the roadway surrounding rock. Accordingly, the roof-to-floor convergence will be smaller, and the damage to the surrounding rock will be slighter. These data indicate that ORFNPT is of a milder roadway surrounding rock deformation than GSEDSCP and is thus more conducive to the maintenance and stability of the lower-section roadway.


## 4.2.1. Stress Distribution Law of Roadway Surrounding Rock in Lower-Section Tunneling

The stress state of roadway surrounding rock is an important indicator to evaluate the stability of roadway, yet the stress characteristics of the coal rock for roadway protection in the gob-side roadway are more crucial for the success of the retained gob-side roadway. Therefore, numerical simulation was conducted on the above two tunneling methods. The yielded stress distribution nephograms is given in Figures5 and 6, and the stress variation curves are displayed in Figure 7. It can be known that under the same geological conditions, for different tunneling methods, the distribution characteristics of the roadway surrounding rock of the excavated roadway differ notably. For ORFNPT, the peak stress of the solid coal side is located at the position of about 3 m in the coal; the peak stress is about 41.5 MPa; the stress concentration coefficient is about 1.74, and the stress on the filling body is about 21 MPa. For GSEDSCP, the stress peak in the solid coal on the lower side of the roadway is about 4.25 m in the coal; the stress peak is around 43.2 MPa; and the stress concentration coefficient is 1.82. Besides, a stress concentration core (range about 1 m and stress peak about 17.2 MPa) exists in the small coal pillar. Compared with GSEDSCP, ORFNPT can protect the roadway tunneling in the lower section more effectively. Its peak stress of the solid coal on the lower side of the roadway is closer to the roadway sides, and the stress of the filling body is much higher than that of the small coal pillar, indicating that the filling body can support the overlying roof better and can bear more load from the overlying strata.Figure 5
Nephogram of vertical stress distribution of surrounding rock in ORFNPT along the coal seam dip.Figure 6
Nephogram of vertical stress distribution of surrounding rock in GSEDSCP along the coal seam dip.Figure 7
Stress distribution curves of roadway surrounding rock along the coal seam dip.


## 4.2.2. Deformation and Displacement Distribution of Roadway Surrounding Rock in Lower-Section Tunneling

The displacement variations of roadway surrounding rock of the two tunneling methods were monitored (Figure8). For ORFNPT, the maximum displacement of the roof is 0.63 m; the maximum heave of the floor is about 0.35 m; and the horizontal displacement on the solid coal side is about 0.51 m. For GSEDSCP, the maximum displacement of the roof is about 1.21 m; the maximum heave of the floor is about 0.48 m; and the maximum horizontal displacement on the solid coal side is about 1.22 m.Figure 8
Distribution characteristics of roadway surrounding rock displacement in ORFNPT and GSEDSCP along the coal seam dip.
(a)
Variation curves of vertical displacement(b)
Variation curves of horizontal displacementThe comparison in Figure7 suggests that the stress field and displacement field of roadway surrounding rock differ under different strength of the filling body. Under a high strength of the filling body, the stress is more likely to be concentrated on the filling body and less likely to be concentrated on the roadway surrounding rock. Accordingly, the roof-to-floor convergence will be smaller, and the damage to the surrounding rock will be slighter. These data indicate that ORFNPT is of a milder roadway surrounding rock deformation than GSEDSCP and is thus more conducive to the maintenance and stability of the lower-section roadway.


## 5. Engineering Application and Effect

### 5.1. Design of Roadway Support Parameters [31]

The 62310 return airway, in which no coal pillars are retained, is a newly excavated roadway along the filling body in the upper section. Considering the influence of roadway section and coal seam dip angle, the section of the 62310 return airway is designed into an inclined rectangular section. Such a design could effectively avoid roof instability in the upper and lower roadway clamping area. The specific support design is given in Figure9.
(1)
The inclined roof was high-strength left-handed special deformed steel bolt without longitudinal reinforcement (ϕ22mm×2500mm), with 7 bolts per row and spacing of 800mm×1,600mm. The first bolt on the high side of the inclined roof was arranged at 30° with the normal direction of the roof; the adjacent bolt was arranged at 20° with the normal direction of the roof; and other bolts were arranged perpendicular to the inclined roof. Three Z2560 new full-length resin anchoring agents were used for full-length anchoring. Anchor cables were high-strength low-relaxation prestressed steel strands (ϕ22 mm, 1×19 strands); the length was 6,500 mm; anchor cables were arranged by 7-0-7 and spacing of 800×1,600mm. The first anchor cable near the high side of the roof was arranged at 30° with the normal direction of the roof; the adjacent anchor cable was arranged at 20° with the normal direction of the roof; other anchor cables were arranged perpendicular to the roof, and the anchor cables were installed on the W4-280 steel bands and staggered with bolts. 4,900 mm long W4-280 steel bands were used for roof protection(2)
The high-strength left-handed special deformed steel bolt without longitudinal reinforcement (ϕ22mm×2,500mm) was used in the side. Six anchors were used in the high side, and the spacing was 800mm×800mm. The anchor near the floor was arranged at 30° in the horizontal direction; other bolts were set perpendicular to the roadway sides. Three bolts were used in the low side, and the spacing was 800mm×800mm. The bolts close to the floor and roof were arranged at 30° in the horizontal direction. W4-280 steel bands were used for roof protection. The length of the low side was 1,800 mm, and the high side was composed of two strips, one being 1800 mm long and the other being 2,250 mm long. Reinforcement support was applied to the high side in the following way: two hollow grouting anchors (ϕ22 mm and length 3,300 mm) were arranged in every two rows of anchors (arrangement mode 2-0-2). The anchor cable near the floor was downward arranged at 30° with horizontal direction, and the other anchor cable was arranged in the middle of bolt 3 and bolt 4Figure 9
Support design in ORFNPT.


### 5.2. Surface Deformation of the 62310 Return Airway

To further explore the application effect of the ORFNPT technology in the mining process of the B10 coal seam in Xinzhuangzi Coal Mine, roof-to-floor convergence and side-to-side displacement of the 62310 return airway were monitored continuously during both tunneling and recovery. According to the monitoring results, during tunneling along the original roadway filling body in the upper section, the roadway gradually stabilizes at the position of about 27 m behind the excavation working face. Roof-to-floor convergence (about 89 mm) is greater than side-to-side displacement (about 58 mm), as illustrated in Figure10. The on-site application effect is shown in Figure 11. During the recovery of the 62310 working face, roof-to-floor convergence and side-to-side displacement of the return airway start to increase continuously at about 100 m away from the working face, and the deformation rate of roadway surrounding rock reaches the maximum at about 40 m away from the working face. The rates of roof-to-floor convergence and side-to-side displacement for the return airway are about 29 mm/d and 50 mm/d, respectively (Figures 12 and 13), and their maximum amounts are about 910 mm and 1,290 mm, respectively (due to the continuous deformation of the roadway, floor dinting and flitching were carried out on the roadway. The deformation of the recovery roadway includes the amount of roadway repair, in which the floor dinting amount is about 500-600 mm and the flitching amount is about 300-400 mm). Clearly, the full-cycle displacement rates and amounts of the surrounding rock of the 62310 return airway satisfy the requirements for stability control of the coal mine roadway, indicating that the ORFNPT technology for stope roadway layout meets the needs for safe mining and realizes the continuous replacement of the stope without coal pillars.Figure 10
Variation curves of surface displacement in the 62310 return airway during tunneling.Figure 11
Support effect of the high side in the 62310 return airway.Figure 12
Variation curves of surface displacement amount in the 62310 return airway during recovery.Figure 13
Variation curves of surface displacement rate in the 62310 return airway during recovery.


## 5.1. Design of Roadway Support Parameters [31]

The 62310 return airway, in which no coal pillars are retained, is a newly excavated roadway along the filling body in the upper section. Considering the influence of roadway section and coal seam dip angle, the section of the 62310 return airway is designed into an inclined rectangular section. Such a design could effectively avoid roof instability in the upper and lower roadway clamping area. The specific support design is given in Figure9.
(1)
The inclined roof was high-strength left-handed special deformed steel bolt without longitudinal reinforcement (ϕ22mm×2500mm), with 7 bolts per row and spacing of 800mm×1,600mm. The first bolt on the high side of the inclined roof was arranged at 30° with the normal direction of the roof; the adjacent bolt was arranged at 20° with the normal direction of the roof; and other bolts were arranged perpendicular to the inclined roof. Three Z2560 new full-length resin anchoring agents were used for full-length anchoring. Anchor cables were high-strength low-relaxation prestressed steel strands (ϕ22 mm, 1×19 strands); the length was 6,500 mm; anchor cables were arranged by 7-0-7 and spacing of 800×1,600mm. The first anchor cable near the high side of the roof was arranged at 30° with the normal direction of the roof; the adjacent anchor cable was arranged at 20° with the normal direction of the roof; other anchor cables were arranged perpendicular to the roof, and the anchor cables were installed on the W4-280 steel bands and staggered with bolts. 4,900 mm long W4-280 steel bands were used for roof protection(2)
The high-strength left-handed special deformed steel bolt without longitudinal reinforcement (ϕ22mm×2,500mm) was used in the side. Six anchors were used in the high side, and the spacing was 800mm×800mm. The anchor near the floor was arranged at 30° in the horizontal direction; other bolts were set perpendicular to the roadway sides. Three bolts were used in the low side, and the spacing was 800mm×800mm. The bolts close to the floor and roof were arranged at 30° in the horizontal direction. W4-280 steel bands were used for roof protection. The length of the low side was 1,800 mm, and the high side was composed of two strips, one being 1800 mm long and the other being 2,250 mm long. Reinforcement support was applied to the high side in the following way: two hollow grouting anchors (ϕ22 mm and length 3,300 mm) were arranged in every two rows of anchors (arrangement mode 2-0-2). The anchor cable near the floor was downward arranged at 30° with horizontal direction, and the other anchor cable was arranged in the middle of bolt 3 and bolt 4Figure 9
Support design in ORFNPT.


## 5.2. Surface Deformation of the 62310 Return Airway

To further explore the application effect of the ORFNPT technology in the mining process of the B10 coal seam in Xinzhuangzi Coal Mine, roof-to-floor convergence and side-to-side displacement of the 62310 return airway were monitored continuously during both tunneling and recovery. According to the monitoring results, during tunneling along the original roadway filling body in the upper section, the roadway gradually stabilizes at the position of about 27 m behind the excavation working face. Roof-to-floor convergence (about 89 mm) is greater than side-to-side displacement (about 58 mm), as illustrated in Figure10. The on-site application effect is shown in Figure 11. During the recovery of the 62310 working face, roof-to-floor convergence and side-to-side displacement of the return airway start to increase continuously at about 100 m away from the working face, and the deformation rate of roadway surrounding rock reaches the maximum at about 40 m away from the working face. The rates of roof-to-floor convergence and side-to-side displacement for the return airway are about 29 mm/d and 50 mm/d, respectively (Figures 12 and 13), and their maximum amounts are about 910 mm and 1,290 mm, respectively (due to the continuous deformation of the roadway, floor dinting and flitching were carried out on the roadway. The deformation of the recovery roadway includes the amount of roadway repair, in which the floor dinting amount is about 500-600 mm and the flitching amount is about 300-400 mm). Clearly, the full-cycle displacement rates and amounts of the surrounding rock of the 62310 return airway satisfy the requirements for stability control of the coal mine roadway, indicating that the ORFNPT technology for stope roadway layout meets the needs for safe mining and realizes the continuous replacement of the stope without coal pillars.Figure 10
Variation curves of surface displacement in the 62310 return airway during tunneling.Figure 11
Support effect of the high side in the 62310 return airway.Figure 12
Variation curves of surface displacement amount in the 62310 return airway during recovery.Figure 13
Variation curves of surface displacement rate in the 62310 return airway during recovery.


## 6. Conclusions

Difficulty in continuous mining replacement caused by complex layout of roadways and unreasonable layout of stope is a significant problem that perplexes coal mining. Aiming at solving the problem, this study analyzed the stress characteristics and design parameters of the original roadway filling body based on the theoretical study on gob-side roadway and verified the stability of roadway surrounding rock in ORFNPT through numerical simulation and on-site monitoring of roadway surrounding rock deformation. The main conclusions are as follows:(1)
The superimposed continuous layer theory was adopted to establish the stress model of the original roadway filling body and obtain the expression of the working resistance of the original roadway filling body. The required strength for the filling body is primarily determined by two factors, i.e., the span of the hanging roof that lies over the filling body and the width of the filling body. The span of the hanging roof is positively correlated with the required strength of the filling body. However, when the width of the filling body reaches a certain value, its further increase fails to change the required strength of the filling body(2)
The finite element calculation software FLAC3D was used to comparatively analyze the stability of roadway in ORFNPT and GSEDSCP. The stress distribution characteristics of the surrounding rock of the roadway are obviously different. When ORFNPT is applied to the lower-section roadway, the peak stress position in the solid coal on the lower side of the roadway is closer to the roadway sides, and the filling body is of a much higher stress than the small coal pillars. Besides, the roadway surrounding rock undergoes milder deformation, and the roadway excavated in the lower section is more stable(3)
As the 62210 transportation roadway of Xinzhuangzi Coal Mine belongs to a straight wall semicircular section, the 62310 return airway adopts an inclined roof trapezoidal section. The hollow grouting anchor cable is used to strengthen the support for the higher side of the roadway. In this way, surrounding rock deformation in ORFNPT is controlled. According to the on-site monitored data, roadway surrounding rock deformation in the 62310 return airway is well controlled during both tunneling and recovery



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*Source: 2545442-2023-03-13.xml*