The Stress Evolution of Adjacent Working Faces Passing through an Abandoned Roadway and the Damage Depth of the Floor

Abstract

An advanced break or a vast region of pressurisation may occur when the working face passes through an abandoned roadway, resulting in a roof falling or water inrush. The stress evolution of the working face passing through an abandoned roadway in a coal mine was comprehensively discussed using theoretical analysis, numerical simulation, and field monitoring. In this study, the calculated critical width of the abandoned roadway where the main roof was bound to an advanced break was 5.4 m. Reducing the suspended length of the main roof is beneficial to the stability of the working face’s passage across the abandoned roadway. The maximum abutment stress on the roof occurred at the working face through a semi-abandoned roadway, reaching 44.3 MPa. Subsequently, it decreased sharply until the working face completely passed and returned to the normal level. The damage depths of the floor strata from the field monitoring were 15 and 20 m, which showed that the use of hydraulic fracturing technology combined with floor grouting and hydraulic support for the abandoned roadway was proposed to stabilise the working face for safe mining.

1. Introduction

The coalfields in northern China are an important coal mining base and mining in this region is often affected by Ordovician limestone aquifer, which hides the safety risk of major water-inrush accidents [1]. However, as most of China’s traditional coal industry adopted a relatively crude operation method for mining, the resource utilisation rate remained low, and many old and abandoned roadways have been left as a result of disorderly mining, which has caused great difficulties in the recovery of working faces [2]. Consequently, mining is plagued by the working face passage through the front roadway. If the main roof of the coal seam breaks in advance when the working face passes through an abandoned roadway, the length of the primary fracture of the basic roof will be greater than that of the normal working face. The main roof may have an advanced or longer break than normal when the working face passes through the forward abandoned roadway [3,4]. This could result in a vast region of pressurisation, which could cause roof falling or coal wall spalling in less serious situations, or it could deepen the damage to the floor strata and result in a water-inrush accident.

Many studies have investigated the stability of roofs and working faces through abandoned roadways. Strength degradation is believed to be the root cause of fracturing of the main roof, which eventually leads to collapse [5,6,7]. Bai et al. suggested that roof collapse can be controlled by changing the width of coal pillars and the stopping time of excavation [8]. Coggan et al. suggested that the thickness of the weak layer of the roof has a significant impact on the degree of damage, and this part needs to be reinforced [9]. Cheng and Jiang believed that the width of the coal pillar is the main factor affecting the stability of the roadway roof. Under in situ stress, structural integrity can be maintained through appropriate support [10,11]. Yu et al. analysed the differences in the stress and plastic zones of different support schemes and proposed a method of using long and short bolts for collaborative support, which can significantly improve the stability of the roadway [12]. Yuan et al. revealed the interaction mechanism of rock-bolt structures in large deformation roadways and increased the thickness of bolts to improve their compression and shear resistance [13]. Jing et al. examined numerous numerical simulation instances and concluded that numerical simulations are crucial for understanding the fundamentals of rock mechanics and the design of rock engineering systems [14]. Ma et al. improved the reinforcing performance of rock bolts by examining the interaction between the rock bolts and rock mass [15]. Xuesheng et al. [16] concluded that an increase in the working face length and mining thickness aggravates the degree of stress concentration, resulting in an increase in the height of the water-flowing fracture zone of the floor. Bai, Jiang, and Zhang et al. studied various support methods such as high-water materials, wooden piles, pumping pillar supporting, backfilling gangue, and anchor cables, and believed that the pre-filling and support method for the abandoned roadway could effectively reduce the stress concentration of the surrounding rock, and is an effective method to prevent the roof from an advanced break [17,18,19,20]. Pan et al. deduced the critical width of coal pillar instability using the key block theory [21]. Luo et al. believed that the existence of an abandoned roadway intensifies the joint action of water pressure, mining disturbances, and flooding, which is the main cause of water inrush accidents [22]. Yasitli et al. concluded that the uniform fracturing of the coal seam roof, which could form cracks on the roof, can maintain the uniform collapse of the roof and improve coal mining efficiency [23]. Wang et al. studied the influence of large-section abandoned roadways and the application of high-water-filling materials [24]. Ju et al. studied the influence of the coal pillar depth and interlayer thickness on the coal seam roof collapse and believed that the rotation and movement of the key blocks in the roof above the coal pillar might cause the overlying stress to concentrate on the coal pillar, resulting in coal pillar instability [25]. Esterhuizen et al. verified the feasibility of an entrance support system by monitoring the displacement and stress of a rock mass [26]. Pan et al. studied the stress state of short-distance coal mining and concluded that when mining under a goaf, the weighting step and abutment stress were lower than the normal level [27]. Sun et al. systematically studied the deformation mechanism and support technology of underground roadways and proved and improved the application of stress relief technique and pre-grouting technique underground [28,29,30].

In summary, most of these studies on the stability of the working face passing through the abandoned roadway are focused on a pre-filling and supportive manner of reducing the stress concentration at the abandoned roadway and restoring the original stress state of the coal seam. However, it is rare enough to support the roadway alone in engineering cases, owing to multiple influences, including groundwater, the small distance to other mining fields, and the layout of diagonally intersecting roadways. Additionally, most studies are based on underground mine pressure data, ignoring the influence of the depth of floor damage, which may cause floor water-inrush accidents. To comprehensively understand the stability of the working face through the abandoned roadway, both theoretical analysis and numerical simulations were utilised to model the stress evolution. The reliability of the calculated results was verified by field monitoring of the damage depth of the floor strata. Based on the results, comprehensive measures, including hydraulic fracturing technology to produce active roof release, auxiliary floor grouting reinforcement, and hydraulic support for the abandoned roadway, are proposed to improve the stability of the working face.

2. Project Overview

The Dongpang Mine is part of the Xingtai mining area of the coalfield in Xingtai City, Hebei Province, northern China. The main tectonic lines in this mine area are primarily east–west, north–east, and north–south. The Ordovician, Cambrian, Metasedimentary Great Wall System, and Taikoo Zanhuang Group are the underlying strata of the coal seam, with the Cambrian and Ordovician strata having a greater influence on coal mining. The No. 9119 working face at Dongpang Mine has an underground elevation of −220 to −274 m, leaving a 6 m net coal pillar in the middle of the No. 9118 working face that was previously recovered. An abandoned roadway is located between the track roadway and the belt roadway along the No. 9 coal seam. The sections of the roadway measure 4 m in width and 3 m in height. The layout of the working faces and the location of the abandoned roadway is shown in Figure 1.

The Daqing limestone aquifer in the roof strata, Benxi limestone aquifer, and Ordovician limestone aquifer in the floor strata constitute most of the water source for the No. 9119 working face. The water from the Daqing limestone aquifer is released along the No. 9118 mining field and roadways, posing no threat to this working face. From the base of the No. 9 coal seam, the heights of the Benxi limestone aquifer and the Ordovician limestone aquifer are approximately 13.39–33.38 m (average 22.86 m) and 30.78–54.57 m (average 43.92 m), respectively. The failure zone on the floor of the No. 9 coal seam serves as the primary conduit for water from the Benxi limestone aquifer. However, it can be employed as an early detection and release method because it has a low water head and a steady amount of water. The greatest danger to this mining operation comes from the irregular water-rich and somewhat high-head Ordovician limestone aquifer. A geological histogram of the No. 9119 working face is shown in Figure 2.

3. Stress Evolution of the Working Face Passing through an Abandoned Roadway

With roadway excavation and working face mining, the in situ stress will be redistributed, so the coal walls on both sides of the abandoned roadway and in front of the working face will produce abutment stresses of different widths, $W_1$ and $W_2$. When $W_1$ and $W_2$ do not overlap, the working face is mined normally, preventing the coal pillar between the working face and the abandoned roadway from becoming unstable. With the continuous advancement of the working face, the coal pillar width gradually decreases, and the load above the pillar continues to increase, leading to a width where the coal pillar is bound to lose stability. This is known as the critical width of the coal pillar instability ($W^{*}$). There are three types of abutment stresses which can be distinguished by the width of the coal pillar ($W$).

(1) When $W\ge W_1+W_2$, the working face is not impacted by the abandoned roadway; therefore, the abutment stress can be divided into the raised stress on each side and the original rock stress in the middle. The abutment stress state, which resembles an asymmetric saddle type, is shown in Figure 3. The peak value and width of abutment stress are greater on the goaf-side because the suspended length on the goaf-side is larger than that on the roadway-side. When $W>W_1+W_2$, original rock stress exists in the coal pillar. When $W=W_1+W_2$, the width of the original rock stress is exactly zero, and at that time, the abutment stresses on the goaf-side and roadway-side start to intersect. In this situation, the coal pillar has a good bearing capacity. The roof of the abandoned roadway can be approximated as a beam with a fixed support at both ends and an even load above, and the roof of the mining field can be approximated as a cantilever beam with an even load above. When the suspended length of the main roof ($L_x$) is greater than the periodic weighting step ($l$), that is, $L_x\ge l$, the cantilever beam cannot carry the top load, resulting in the fracture of the main roof at the fixed end. The periodic break line, in this instance, moves forward to the working face via a periodic weighting step.

(2) When $W^{*}<W<W_1+W_2$, the evolution of the abutment stress changes from the saddle-type to the oblique-table type, as indicated in Figure 4, because of the overlap of the abutment stresses on the goaf- and roadway-side. The coal pillar is stable because $W$ is greater than $W^{*}$, which means that the working face does not have an advanced break.

(3) When $W\le W^{*}$, the distribution of abutment stress is presented as a solitary peak. The maximum abutment stress value is obtained when $W=W^{*}$. If the working face advances further, the middle coal pillar will remain in a state of plastic flow for a considerable amount of time, which will cause serious damage and deformation, and the strength and bearing capacity will be significantly reduced, resulting in instability and damage. At this moment, the suspended length of the main roof ($L_x$) increases from $D$ to $A+W+D$, that is, $L_x=A+W+D$, which occurs in two ways. When $L_x=A+W+D<l$, the fixed end of the cantilever beam can support both its own weight and the upper loads without breaking. When $L_x=A+W+D\ge l$, the cantilever beam fractures at the fixed end, and the main roof undergoes an advanced break, creating a vast region of pressurisation that might seriously threaten coal mining safety, as shown in Figure 5.

Based on the Bieniawski formula [31]:

$${\sigma }_p=0.235{\sigma }_c(0.64+0.36\frac{W}{M})$$

(1)

where ${\sigma }_p$ is the ultimate compressive strength of the coal pillar (MPa), $W$ is the width of the coal pillar between the working face and abandoned roadway (m), ${\sigma }_c$ is the uniaxial compressive strength of the standard sample of the coal pillar (MPa), and $M$ is the height of the coal pillar (m).

The coal pillar is not only required to carry the load of its own overlying strata but also the load of the overlying strata of one-half of the abandoned roadway and a significant portion of the load of the overlying strata of the cantilever beam on the goaf-side, so the static load set on the coal pillar is:

$$q=\gamma H(\frac{W+\frac{A}{2}+kD}{W})$$

(2)

where $q$ is the static load setting on the coal pillar, $\gamma$ is the average volume weight of the overlying rocks (kN/m³), $H$ is the average thickness of the overlying strata (m), $A$ is the width of the abandoned roadway (m), $k$ is the stress concentration factor (k = 1.5~5.0), and $D$ is the length of the cantilever beam (m).

When the ultimate compressive strength of the coal pillar is equal to its static load, that is ${\sigma }_p=q$, coupling (1) to (3) results in

$$0.235{\sigma }_c(0.64+0.36\frac{W}{M})=\gamma H(1+\frac{\frac{A}{2}+kD}{W}).$$

(3)

The critical width of coal pillar instability can be derived from Equation (4):

$$W^{*}=\frac{-b+\sqrt{b^2-4ac}}{2a}$$

(4)

where $a=\frac{0.085{\sigma }_c}{M}, b=0.15{\sigma }_c-\gamma H$, and $c=-\gamma H(\frac{A}{2}+kD)$.

If the width of the coal pillar is less than the critical width, an advanced break occurs in the main roof. Because the suspended length on the goaf-side ($D$) is uncertain, the static load applied to the coal pillar is the smallest when $D$ is 0, and at that time, the critical width of the abandoned roadway, where the main roof will inevitably experience an advanced break, can be deduced.

$$A^{*}=l-W^{*}-D=l-\frac{-b+\sqrt{b^2-4ac}}{2a}.$$

(5)

According to the field data, the periodic weighting step of the No. 9119 working face was 20 m, the height of the coal pillar was 6 m, the uniaxial compressive strength of the No. 9 coal seam was approximately 20 MPa, the overlying strata was 250 m from the ground level, and its average volume weight was 24,000 kg/m³. Using Equation (5) and the above data, we can obtain $A^{*}$ as 5.4 m. Therefore, an advanced break in the main roof will undoubtedly occur when the width of the abandoned roadway is greater than 5.4 m. Because the width of the abandoned roadway in this mining area is 4 m, which is less than the necessary width, an advanced break may not occur in the main roof.

According to $L_x=A+W+D$, the width of the coal pillar, suspended length of the main roof, and width of the abandoned roadway are all significant factors in determining whether the main roof will advance break. $W^{*}$ is used as a rough alternative to $W$ because the strength of the coal pillar rapidly declines when its width is below the critical width. $W^{*}$ is positively correlated with $D$ in Equation (4). Therefore, after the width of the abandoned roadway is established, there is a positive association between $L_x$ and $D$. For the working face to pass through the abandoned roadway safely, it is crucial to reduce the suspended length of the main roof.

4. Numerical Simulation

Theoretical analysis is often based on idealised models and necessary assumptions. Considering the complexity of the project, a 3D numerical model created with Flac^3D software can more accurately depict the dynamic evolution of the No. 9119 working face passing through the abandoned roadway.

4.1. Three-Dimensional Numerical Model

A 3D model with an x-axis of 302 m, y-axis of 180 m, and z-axis of 80 m was created, as shown in Figure 6a. The working face was 90 m long. Each cross-section of the track, belt, and abandoned roadways measured 4 m in width and 3 m in height, with a 53° angle between them. Given the lateral abutment stress, two coal pillars of 50 m width were set aside on either side of the No. 9118 and No. 9119 working faces. The main roof fractured periodically owing to the large distance of the open cut from the roadway intersection, which allowed the model to be excavated from the boundary section. The bottom of the model was fixed, and the horizontal movement of its four sides was constrained. The upper surface of the model, whose boundary conditions are schematically depicted in Figure 6b, was subjected to an even load of $P=\gamma h=6.0 \mathrm{MPa}$.

The physical and mechanical parameters of the rock formation at the working face based on drilling data collected on-site are listed in

Table 1. Physical and mechanical parameters of rock formation.

Rock Formation	Thickness (m)	Bulk Modulus (GPa)	Shear Modulus (GPa)	Cohesion (MPa)	Internal Friction Angle (°)	Tensile Strength (MPa)	Density (kg/m3)
Siltstone	6	13.7	9.4	6.0	35	1.5	2400
Fine-sandstone	11	17.9	11.8	4.9	36	2.5	2500
Daqing limestone	5	25.6	19.2	9.5	40	4.9	2700
No. 8 Coal	2	4.8	1.6	0.3	28	0.5	1400
Siltstone	6	12.8	8.1	3.8	32	1.0	2400
No. 9 Coal	6	4.8	1.6	0.3	28	0.5	1400
Carbon mudstone	3	11.1	5.4	2.6	29	0.8	2500
Mudstone	6	9.8	5.1	2.7	30	0.5	2500
Medium grained sandstone	14	8.9	6.1	3.6	29	1.8	2500
Benxi limestone	3	12.6	9.1	5.9	36	3.6	2800
Fine-sandstone	6	25.8	13.3	5.2	36	2.0	2400
Bauxite mudstone	12	15.7	6.0	1.7	28	0.5	2500

.

The fallen gravel in the mining field is a kind of loose medium, but Flac^3D belongs to the continuous medium simulation method. Therefore, we approximate the supporting effect of the gravel on the roof to elastic support [32]. The gravel will gradually be compacted by the overlying strata; therefore, the physical and mechanical parameters are positively correlated with time. Based on the physical and mechanical parameters of the rock formation and the actual mining pace, a demarcation line 40 m behind the working face is advised. The deformation parameters of gravel in the mining field, as determined using empirical formulas [32], are listed in

Table 2. Physical and mechanical parameters of gravel in the mining field.

Distance from the Working Face (m)	Bulking Coefficient	Elastic Modulus (MPa)	Poisson Ratio	Bulk Modulus (MPa)	Shear Modulus (MPa)	Density (kg/m3)
<40	1.5	26.6	0.06	10.1	12.5	1650
>40	1.3	37.4	0.08	14.8	17.3	1700
NO. 9118 mining field	1.2	189.0	0.25	126.0	75.6	2400

.

$$\rho =1600+800\left(1-e^{-1.25t}\right),$$

(6)

$$E=15+175\left(1-e^{-1.25t}\right),$$

(7)

$$\mu =0.05+0.2\left(1-e^{-1.25t}\right),$$

(8)

where $t$ is the time after the collapse of the immediate roof of the mining field (a).

In order to ensure that the roof elements would not invade the floor elements after the excavation, one contact surface with very large physical and mechanical properties was set on the floor of the coal seam. The stiffness-normal was 100 GPa/m, stiffness-shear was 100 GPa/m, tensile strength was 10 GPa, internal friction angle was 30°, cohesion was 10 GPa, and dilation was 6°.

4.2. Results

4.2.1. Initial Stress

The vertical stress cloud of the initial ground stress measured in Pa is shown in Figure 7. The following clouds and diagrams units are Pa unless otherwise indicated. Before the mining field was excavated, the initial ground stress reached an equilibrium. From there, the stress increased consistently with depth, peaking at 7.8 MPa at the bottom of the model.

The vertical stress cloud after the No. 9118 mining field was excavated is depicted in Figure 8. High compressive stress was produced by the two coal walls on either side of the No. 9118 mining field, which was partially concentrated in the No. 9119 working face, causing significant difficulties in the mining of the No. 9119 working face.

4.2.2. Stress Evolution of the Roof

When the periodic weighting step is 10 m, 3D diagrams of the vertical stress at 1 m above the roof for excavations of 30, 60, 90, and 120 m are shown in Figure 9.

As the No. 9119 working face advanced by 30 m, the peak vertical stress on the roof, as can be seen in Figure 9a, occurred in front of the No. 9119 working face offset towards the No. 9118 mining field, reaching 34.0 MPa, with a stress concentration factor of 5.2. Only a 6 m wide coal pillar was reserved between the two working faces, which was insufficient to compensate for the lateral abutment stress and caused damage. As a result, there was a significant amount of stress concentration because the lateral abutment stress from the No. 9118 mining field was transferred to the inner side of the No. 9119 working face and superimposed on the front abutment stress.

The front abutment stress reached 36.9 MPa, and the stress concentration factor reached 5.7 when the working face advanced 60 m. At that point, the working face started to expose the abandoned roadway, as illustrated in Figure 9b. The No. 9118 mined-out area was compacted once more with the extraction of the No. 9119 working face. The working face passed half of the abandoned roadway at a distance of 90 m, where the abutment stress on the roof of the remaining triangular coal pillar reached its peak value of 44.3 MPa, as shown in Figure 9c. When the working face was entirely through the abandoned roadway, the peak abutment stress on the roof, as shown in Figure 9d, was 30.5 MPa.

The relationship between the peak vertical stress 1 m above the roof and the mining position of the working face is shown in Figure 10. The peak stress increased slowly when the working face was farther from the abandoned roadway. The area of the triangular coal pillar gradually decreased as the working face continued to advance beyond 60 m, and the peak stress on the roof began to increase sharply. The maximum stress occurred at 90 m, where the working face passed through a semi-abandoned roadway, where the maximum stress reached 44.3 MPa. This was because the abandoned roadway cut off the coal body behind it and the triangular coal pillar, resulting in the superposition of the front abutment stress, the lateral abutment stress of the No. 9118 mining field, and the abutment stress of the abandoned roadway, which caused a large degree of stress concentration. The triangular coal pillar area decreased as the working face advanced, and as a result, it could no longer support the roof. Most of the abutment stress began to transfer to the coal body behind the abandoned roadway, which led to a more uniform distribution of the abutment stress. As a result, the peak abutment stress began to decrease until the working face completely passed and returned to the normal level.

Side views of the vertical stress cloud of the selected excavation are shown in Figure 11. Vertical stress appears in the sequence of the saddle type, oblique table type, and solitary peak type with the excavation of the working face.

(1)

As shown in Figure 11a, when the width of the coal pillar was 92 m in the middle of the working face and the abandoned roadway, the vertical stress of the coal pillar exhibited an asymmetric saddle type, with the peak stress (16.3 MPa) on the goaf-side being larger. The coal pillar midsection still contained the original rock stress zone, indicating that the front abutment stress was not affected by the abandoned roadway.

(2)

As shown in Figure 11b, when the width of the coal pillar was 52 m, the vertical stress of the coal pillar was an oblique table type, and the front abutment stress increased significantly, with peak stress of 30.1 MPa. The abutment stresses on both sides influenced one another.

(3)

As shown in Figure 11c, when the width of the coal pillar was 32 m, the vertical stress of the coal pillar was an asymmetrical peak type, with peak stress of 33.9 MPa. Currently, the abutment stress of the central coal pillar affects the abutment stress of the rear coal wall of the abandoned roadway, which results in a significant increase in the latter.

4.2.3. Effect of Weighting Steps on the Stability of the Roof and Floor

A common multiple of 60 m excavation length was selected for comparative analysis using Flac^3D to simulate the weighting steps of 10, 20, and 30 m, respectively. Figure 12 depict the vertical stress clouds of the coal pillar at 0.5 m above the floor before the roof collapses. When the weighting steps are 10, 20, and 30 m, the peak vertical stresses of the coal pillar are 43.8, 45.5, and 45.9 MPa, respectively, which indicates that reducing the weighting step can reduce the maximum front abutment stress on the coal pillar. Consequently, active roof cutting can be used to control the suspended length of the main roof behind the working face, thereby reducing the front abutment stress.

From the above analysis, the largest front abutment stress occurred in front of the No. 9119 working face offset toward the No. 9118 mining field; therefore, the side views of the plastic zone near the No. 9118 mining field at 120 m of excavation were chosen, as shown in Figure 13. Shear failure primarily occurred in the plastic zone in the stope floor strata. Shear-n indicates that shear failure is occurring in the elements in the current cycle, while Shear-p indicates that shear failure has occurred in the elements in the previous cycle. Therefore, increasing the weighting step would cause the shear failure of the elements in front of the working face faster. The maximum depth of the plastic zone in the stope floor strata was 23 m when the weighing step was 10 m. The major portion of the plastic zone did not change significantly when the weighted step was increased to 20 and 30 m, but new plastic zones of various sizes appeared from 32 to 36 m below the floor, with the largest plastic zone occurring at 30 m. With the increase of the weighting step distance, the floor failure depth increased firstly and then tended to be stable [33]. Therefore, when the weighting step increased from 20 m to 30 m, the depth of the plastic zone did not change significantly. Combined with the physical and mechanical parameters of the floor strata, no plastic zone has arisen in the range of 23 m to 32 m below the floor because of the Benxi limestone and fine sandstone in this area, which are relatively hard in texture and have large shear and tensile strengths. A certain zone of plasticity developed because of the lithological weakness of the bauxite mudstone that lies beneath the fine sandstone, which could still be inferred from the fact that as the weighting step increased, the extent of the plastic zone in the stope floor strata increased, indicating an increase in the depth of floor failure. However, the floor failure depth would not increase indefinitely and would eventually become stable.

5. Results

According to theoretical analysis and numerical simulation, the maximum front abutment stress of the coal pillar and roof can be reduced, and the depth of the plastic zone in the stope floor strata can be reduced by actively reducing the suspended length of the main roof. The Ordovician limestone aquifer had an impact on the No. 9119 working face, possibly causing water damage, so the floor strata could be grouted to strengthen the mechanical properties of the rock. Simultaneously, hydraulic supports were added to the interior of the abandoned roadway to ensure the safe passage of the No. 9119 working face.

5.1. Application of Hydraulic Fracturing Technology

The No. 9119 working face used hydraulic fracturing technology to actively cut the roof strata and control the weighting step to be 10 m. Holes were drilled in twin groups A and B. Group A was drilled from the top of the belt roadway to the interior of the roof strata of the mining field, with a length of 50 m, an elevation angle of 45°, and hole spacing of 10 m. It was used to cut the link between the main roof and the rock above and release the pressure. Group B was drilled vertically upward in the middle of the top of the belt roadway with a length of 34 m and a hole spacing of 10 m. The purpose of Group B was to cut the connection between the No. 9119 and No. 9118 mining fields and to reduce the impact of the No. 9118 mining field. Group A and group B holes were arranged in a staggered manner and extended close to the stopping line. Layout diagrams of the holes are shown in Figure 14.

5.2. Floor Grouting Reinforcement

Underground structures are subject to weathering (for example, by water), which can reduce their mechanical strength and may eventually lead to unstable excavations [34,35], so the use of floor grouting reinforcement technology can significantly reduce the impact of groundwater on the strength of the floor strata. To achieve the best reinforcement effect and to consider the economics of the project, floor strata from 10 m below the floor to the top of the Ordovician limestone aquifer were used as the target rock strata for grouting. The key grouting reinforcement areas include the physical prospecting of unusual areas, typical hydrogeological areas, and areas where the overtopped hole has not been explored. A total of 360 advancing and grouting holes were constructed in the No. 9119 working face, with 23,895.5 m drill footage and 246.71 t grout.

5.3. Support for the Abandoned Roadway

The section size of the abandoned roadway at the No. 9119 working face is 4.2 m wide at the top, 4.5 m wide at the bottom, and 2.8 m high at the middle. DW-28 hydraulic supports with DJB (S) metal-hinged beams were used to strengthen the abandoned roadway. The hydraulic supports were spaced 0.8 m along the abandoned roadway, 0.3 m from the coal wall on both sides, and 1.3 m between each row of pillars, which provided 0.3 MPa for the abandoned roadway.

5.4. Verification Using the Field Monitoring

To monitor the displacement of the floor strata, vibrating wire multi-point displacement meters with two measuring stations were used at depths of 10, 15, 20, and 25 m below the floor. Station No. 1 was buried in the chamber of the track roadway away from and unaffected by the abandoned roadway, whereas station No. 2 was buried near the intersection of the track roadway and abandoned roadway, where it was affected by the abandoned roadway. The locations of the measuring stations and multi-point displacement meters are shown in Figure 15.

A peephole was used to inspect the borehole before the multi-point displacement gauge was buried. The conditions of borehole No. 1 from the peephole are shown in Figure 16a. A small number of fissures were produced at 5–8 m, which gradually decreased until they disappeared with increasing depth. The conditions of borehole No. 2 from the peephole are shown in Figure 16b. A small amount of water inflow is produced at 6–8 m, where mud occurs owing to the mudstone layer, and vertical fissures are produced at 12–15 m.

The relationship between the vertical displacement of each measuring point and the distance between the working face and station is shown in Figure 17. The vertical displacement of each measuring point decreased and then increased as the working face advanced, which was caused by the front abutment stress being transferred to the floor strata, resulting in compression deformation. As the working face moved forward, the peak front abutment stress passed through the measuring station, which decreased the stress concentration factor above the measuring point. The rock formation started to expand because of the floor strata pressure relief, which resulted in tensile deformation at the measuring point in the direction of the upper free space. The floor strata of the No. 1 measuring station experienced discontinuous displacement changes between 10 and 15 m, whereas the floor strata of the No. 2 measuring station experienced discontinuous changes in displacement between 15 and 20 m. The damage depths of the floor strata, which are 15 and 20 m, respectively, are commonly regarded as the upper limit of the aforementioned interval to ensure the safety of pressure mining. This suggests that the abandoned roadway increased the stress concentration at the No. 2 measuring station, which in turn increased the damage depth and raised the possibility of water inrush of the floor strata. The peak front abutment stress under normal mining conditions was approximately 25 m in front of the working face, as shown in Figure 17a, where the turning point of the vertical displacement at the No. 1 measuring station was located. However, as shown in Figure 17b, the peak front abutment stress was approximately 30 m in front of the working face under No. 2 measuring station conditions. In the numerical simulation, the depths of the plastic zone at the two stations were extracted as 13 and 17 m, which is in error with the measured data by approximately 15% and 13%, respectively.

The final monitoring results are in good agreement with the numerical simulation results, indicating that the application of hydraulic fracturing technology, floor grouting reinforcement, and support of the abandoned roadway can reduce the maximum front abutment stress of the mining field, thereby reducing the damage depth of the floor strata. In this case, the No. 9119 working face passed through the forward abandoned roadway without any advanced break or weighing over a great extent, and there were no water inrush accidents.

6. Conclusions

(1)

The stress evolution of the working face passing through the abandoned roadway can be characterised by the types of saddle, oblique table, and solitary peak. The critical width of the abandoned roadway where the main roof is bound to advanced break was calculated to be 5.4 m. With the width of the abandoned roadway determined, the suspended length of the main roof was the key factor in determining whether the main roof will advance break.

(2)

Using elastomeric expressions to approximate fallen gravel, a Flac^3D simulation method was proposed to simulate the stress evolution laws and the effect of different weighting steps. The error between the depths of the plastic zone obtained using this simulation method and the field monitoring data was within 15%.

(3)

The peak stress on the roof increased slowly when the working face was farther away from the abandoned roadway and then increased sharply when the working face began to expose the abandoned roadway. The maximum stress occurred at the working face through a semi-abandoned roadway, where the maximum stress reached 44.3 MPa. Subsequently, the area of the triangular coal pillar decreased, so the triangular coal pillar could no longer support the roof. The coal wall behind the abandoned roadway carried most of the front abutment stress, which caused the peak stress to slope downward.

(4)

The use of hydraulic fracturing technology to produce active roof release, with the aid of floor grouting reinforcement and support for abandoned roadways, was proposed as a holistic management technique. The damage depths of the floor strata from the field monitoring were 15 and 20 m, respectively, which showed that the measurements enabled the No. 9119 working face at Dongpang Mine to safely pass through the forward abandoned roadway and provided a significant reference for similar projects.