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Article

Analysis of Surrounding Rock Control Technology and Its Application on a Dynamic Pressure Roadway in a Thick Coal Seam

by
Jian Hao
1,2,
Anfa Chen
2,*,
Xuelong Li
1,2,*,
Hua Bian
3,
Guanghua Zhou
4,
Zhenguo Wu
5,
Linjun Peng
6 and
Jianquan Tang
2
1
Mine Disaster Prevention and Control-Ministry of State Key Laboratory Breeding Base, Shandong University of Science and Technology, Qingdao 266590, China
2
College of Energy and Mining Engineering, Shandong University of Science and Technology, Qingdao 266590, China
3
School of Energy and Mining Engineering, China University of Mining & Technology-Beijing, Beijing 100083, China
4
Guoneng Ningxia Coal Industry Group Co., Ltd., Yinchuan 750011, China
5
Ningxia Coal Science and Technology Research Institute Co., Ltd., Yinchuan 750011, China
6
School of Civil Engineering and Architecture, Dalian University, Dalian 116622, China
*
Authors to whom correspondence should be addressed.
Energies 2022, 15(23), 9040; https://doi.org/10.3390/en15239040
Submission received: 26 October 2022 / Revised: 16 November 2022 / Accepted: 26 November 2022 / Published: 29 November 2022
(This article belongs to the Section H: Geo-Energy)
Browse Figures
Versions Notes

Abstract

:
The deformation control of roadways surrounded by rock in the fully mechanized amplification sections of extra-thick coal seams is problematic. To analyze the failure and failure characteristics of a support frame, as well as the deformation and failure processes of the surrounding rock, through theoretical analysis and industrial tests, the deformation and support conditions of a return airway of a fully mechanized caving face in an extra-thick coal seam in the Yangchangwan Coal Mine, in the Ningdong mining, area were examined. Combined with limit equilibrium theory and roadway section size, the width of the coal pillar of the return air roadway at the 130,205 working face was calculated to be 6 m. The layout scheme and implementation parameters of roof blasting pressure relief, coal pillar grouting modification, and bolt (cable) support were designed. Based on the analysis, a “Coal pillar optimization–roof cutting destressing–routing modification–rock bolting” system for surrounding rock control in synergy with the fully enlarged section mining roadway in the extra-thick coal seam was proposed, and the deformation of the surrounding rock was monitored, along with the stress of the support body and the grouting effect on the site. Field experiments show that after the implementation of the surrounding rock control in synergy with the roadway, the maximum subsidence of the top plate was 55 mm, the maximum bottom heave of the bottom plate was 55 mm, the maximum values of the upper and lower side drums were 30 mm and 70 mm, respectively, and the breaking rate of the bolt (cable) and the deformation of the surrounding rock of the roadway was reduced by more than 90% and 70%, respectively. The effective performance of the coal pillar grouting was observed as well. Field practice of the roadway surrounding rock control in the synergy method indicated that rock deformation was effectively controlled, and the successful application of this technology was able to provide reliable technical and theoretical support for the Ningdong mining area and mines with similar conditions.

1. Introduction

Reserves of extra-thick coal seams in the coal-producing countries account for 20–50% of the entire recoverable coal resources in the world [1,2,3,4,5,6,7]. Fully mechanized top-coal mining techniques are generally used for extra-thick coal seams. Compared with other techniques, sub-level caving mining has the characteristics of having a large unit yield and other remarkable benefits. The working face excavated using this method is generally 200–400 m long. Reserves of thick coal seams in China are relatively rich, having a wide but uneven distribution, and are found more abundantly in the west of China than in the east. As the mining of the mine goes deeper, the ground stress level of the stope also gradually increases. After the mining of the working face, it is easy to form stress concentration at the coal pillar position, which increases the width of the required roadway coal pillar, making it difficult to support the roadway in the lower section with severe deformation and to maintain the lower coal mining roadway. Therefore, it has become a major engineering problem to be solved urgently to reduce the waste of resources caused by protecting coal pillars and improve the state of the underground pressure in deep stopes [8,9,10,11,12].
The issue of the large deformation control of surrounding rock in a deep mining roadway has been extensively analyzed [13,14,15,16]. In the early 20th century, classical mechanical techniques began to be introduced into rock mechanical analysis for underground engineering. According to the classical geotechnical stress theory presented by Haim and Rankine, stress acting on the support structure is the weight of the overlying strata, and the difference lies in the lateral pressure coefficient [17]. Protodyakonov proposed the Platts’ falling arch theory, highlighting that, after the excavation of a roadway in the rock formation, a naturally balanced arch will be formed above the roadway [18]. Since the 1950s, support related to underground engineering has been analyzed using the principle of elastic–plastic mechanics. The Kastner formula and the Fencer formula [19,20,21] have been the most widely used in this area. This support theory explains the principle of the joint action of support and surrounding rock through the joint action of “support-surrounding rock”. The early energy support theory proposed by Salamon et al. stated that surrounding rock and supporting roadway structures are mutually influenced, and the support structure maintains a balance between them by absorbing the energy released by surrounding rock during the deformation process [22,23]. The strain control theory of surrounding rock support proposed by Hiroshi and Harusuke reported that, as support strength increases, the deformation of the surrounding rock of the roadway gradually decreases [24,25]. The deformation of the surrounding rock can therefore be controlled within a reasonable range by increasing the support strength. Liagel et al. proposed the theory of maximum horizontal stress [26,27,28,29,30,31]. Here, the horizontal stress of the surrounding rock is generally greater than its vertical stress. According to the stress level, horizontal stress can be divided into maximum stress and minimum stress, which differ by several degrees. It was found that, when a roadway is arranged along the maximum horizontal principal stress, the deformation of the surrounding roadway rock is minimal due to roadway stability not being affected by the maximum horizontal principal stress, but by the minimum principal stress. Therefore, under suitable conditions, the roadway should be adjusted as far as possible according to the direction of the maximum principal stress.
Yu et al. proposed the theory of axial deformation to study the stability of surrounding rock [24]. Herein, the axial ratio change of a roadway plays a decisive role in controlling the failure of the surrounding rock. This theory discussed the relationship between the roadway axial ratio and the stability of the surrounding rock [32,33,34,35,36,37]. They found that rock failure surrounding a roadway is not instantaneous, but rather a gradual mechanical process with time and space effects. In terms of space, damage, failure, and deformation occur from one or several parts of the rock mass, developing into the overall instability of the surrounding rock or supporting body. These initial areas experiencing damage are important areas [38]. Based on this concept, the theory of coupling support for key parts was proposed. Kang et al. proposed the one-time support theory of a highly prestressed bolt and developed the high-pressure split grouting modification technology, upon which the supporting–modification–destressing in synergy technology was formed [39,40]. Ma et al. developed and validated the calculation formula of the deviatoric stress of the surrounding rock of a circular roadway under a non-uniform stress field and the calculation technique of the plastic zone radius by investigating the distribution of deviatoric stress field and plastic zone of the surrounding rock [41]. Liu et al. believed that deep coal and rock mass stability is affected by high ground stress, osmotic pressure, and temperature. Based on this belief, a surrounding rock control method of ultra-high-strength bolt support–grouting consolidation–energy release was proposed [42]. Bai et al. highlighted that the basic premise of surrounding rock control in deep roadways is to improve the self-bearing capacity of the surrounding rock, optimize the stress environment, and adopt a reasonable support technology [43]. Based on the failure characteristics and action mechanism of two side anchors in the deep roadway, Gou et al. established a stability mechanical model under two types of instability conditions for two side anchors in the roadway, and proposed roadway bolt- and anchor-cable-coordinated support technology. Jiang et al. studied the use of highly prestressed anchor bolts and robust anchor cables as surrounding rock control technology for dynamic large deformation failure and anti-scour conditions of deep tunnels with different functional types [44,45,46,47,48]. Based on this, the presplitting pressure relief technology of roof blasting has been continuously developed, especially in the hard roof, effectively reducing the weighting step behind the working face and eliminating the suspended roof on one side of the goaf. Tang et al. established a numerical model of steel pipe strain test and performed it to simulate the expansion of SCDA based on the coupling of finite difference and discrete element methods. The expansion process of a specific type of SCDA can be simulated by adjusting the microparameters according to the experiment [49,50].
Current investigation on rock control technology surrounding a roadway predominantly focuses on the eastern mining areas of China. However, limited investigations have been undertaken in the western mining areas, especially for rock control surrounding mining roadways in extra-thick coal seam large sections utilizing a fully mechanized top-coal caving face.
This study examines the rock control technique surrounding a substantial portion of the mining roadway in an extra-thick coal seam using the 130,205 working face of Yangchangwan Coal Mine in the Ningdong mining area as the research object. Based on proposals to increase the surrounding rock bearing capacity, improve the stress environment, and optimize auxiliary support, a four-in-one surrounding rock collaborative control method for coal pillar size optimization, active pressure relief of roof presplitting blasting, active modification of coal pillar high-pressure splitting grouting, and high-strength prestressed bolt support is proposed. The implementation of this method can effectively control the large deformation of the mining roadway in the ultra-large cross-section with a fully mechanized caving face in extra-thick coal seams. The successful application of this technology will provide a reference for other mining areas in western China with similar geological and mining technical conditions.

2. Project Overview

2.1. Study Area

Yangchangwan Minefield is located in Ciyaobao Town, Lingwu City, Ningxia Hui autonomous region, at the heart of the Lingwu mining area, about 60 km away from Yinchuan City. The mining area is surrounded by sand dunes and, while the ground is affected by erosion, overall fluctuation is small. Surface elevation in the study area is high in the south and low in the north, having an elevation of about 1400 m. The highest point of the minefield is 1465.7 m, located in the southernmost black pimple. The lowest point is 1340 m, located in Shuishigou, along the 13 exploration lines in the north.
The 130,205 working face is located in the eastern part of the Yangchangwan No. 1 Minefield. The surface is covered by dunes, and the terrain is low, gentle, and flat, with little fluctuation. The coal seam is classified as “2 coal”, having a buried depth of about 615 m and a coal-seam thickness of 8.6–9.4 m (average 9.0 m). The 130,205 working face is adjacent to the goaf of the 130,203 working face, wherein a 6 m small coal pillar is reserved with the 130,203 belt roadway. The length of the return air roadway is 2198 m (1600 m small coal pillar section), and the cross-sectional dimension is 5000 × 4000 mm. To successfully work with the goaf, the working face adopts a longwall backward mining method, comprehensive top-coal caving mining technology, and a total caving method. According to the site survey and geological exploration, the layout of the mine and the working face is shown in Figure 1, the distribution of the coal strata and geological description is shown in Table 1, and the coal strata distribution and geological description are shown in Table 2.

2.2. Original Support Scheme and Roadway Status

According to the original support scheme in the operation specification, the roadway section has a trapezoidal shape, and the original bolt cable support scheme is shown in Figure 2. A complete set of 22#-M24-2500 mm BHRB500 rebar anchors is used in the roadway roof, having a row spacing of 900 mm. A complete set of 20#-M22-2300 mm BHRB500 rebar anchors is used in the sidewall, having a row spacing of 800 mm. A round steel strip consisting of ϕ16 mm round steel is used as a roof bolt, which is consistent with that used in the roadway section. The roadway side bolt uses a W280-450-5 mm steel plate. The metal mesh on the roof (6000 × 1000 mm) has a mesh size of 150 × 150 mm. The size of the steel strand anchor cable is ϕ22 × 10,300 mm and the anchor plate is arched with a size of 300 × 300 × 16 mm.
Although this support scheme has been implemented in the roadway, severe deformation still exists, especially during mining. Deformation has manifested as large deformation of the roadway side and floor. Cumulative floor heave was reported to be more than 307 mm, and the shrinkage of the two sides was more than 803 mm. The slurry skin of the coal pillar was widely broken, with a large number of roadway shoulder anchor bolts and anchor cables being broken. Supporting components, such as steel belt tearing and support plate overturning, were invalid. It was reported that 453 anchor cables were broken at the top of the coal pillar (598 m) with a return air roadway of 35 m in the 130,205 working face and 0.76 anchor cables broken along a linear meter.

2.3. Behavior Characteristics of Large Coal Pillars

Deformation and failure characteristics were identified in surrounding rock deformation of the original 35 m coal pillar roadway at the 130,205 working face of the Yangchangwan Coal Mine using an observational technique. According to field observation, the characteristics were as shown in Figure 3:
  • Faults around the roadway mainly include F20, F23, and the five-knot anticline. In this area, the roadway was not only affected by the original gravity stress but it was also affected by residual tectonic stress, especially by an increase in horizontal tectonic stress, resulting in serious damage to the roadway roof and obvious net pocket subsidence.
  • Large initial deformation rate. After tunnel excavation, the initial deformation speed of the surrounding rock was fast, and it was subjected to high stress for a long period of time. Through observation, floor local uplift was seen (Figure 3a). Over time, the deformation of the bottom plate increased, and phenomena such as bending deformation of the steel belt and drilling of the column were observed (Figure 3b).
  • Asymmetrical deformation of roadway surrounding rock. The maximum deformation of the solid-coal surrounding rock was about 633 mm, and the maximum deformation of the coal-pillar-side surrounding rock was about 170 mm. The approaching amount of the solid-coal side was significantly greater than that of the coal-pillar side. On the coal-pillar side, a coal body falling off was identified (Figure 3c). The falling coal body could be directly correlated to abutment pressure distribution in the mine. As the solid-coal side was closer to peak abutment pressure, the stress in the solid coal increased. This resulted in large deformation with coal readily falling off and damage to the coal side. As the side of the coal pillar was far from the peak value of abutment pressure, it was in a stable internal stress field. Moreover, the coal pillar had been reinforced by grouting, and its mechanical properties were enhanced. This resulted in small deformation and small ground pressure behavior strength, which mainly manifested itself as local cracking, and bolt-extrusion failure (Figure 3d).
Energies 15 09040 g003a 550Energies 15 09040 g003b 550
Figure 3. Strata behavior characteristics of large coal pillar roadway protection. (a) Floor heave of roadway; (b) Single pillar drilled into floor; (c) Deformation of both sides of the roadway; (d) Partial cracking of roadway.
Figure 3. Strata behavior characteristics of large coal pillar roadway protection. (a) Floor heave of roadway; (b) Single pillar drilled into floor; (c) Deformation of both sides of the roadway; (d) Partial cracking of roadway.
Energies 15 09040 g003aEnergies 15 09040 g003b

2.4. Failure Principle of the Large Coal Pillar Roadway

2.4.1. Time–Space Effect of Mining Roadway Failure

Space–time conditions of mining roadway excavation have an important impact on the degree of damage on the roadway surrounding rock. Excavating and maintaining the correct position of the roadway over time is the key to ensuring roadway stability and safety. After mining the 130,203 working face, the stress of the surrounding rock at the stope was redistributed. This was divided into three parts according to the level of stress: (i) low-stress area includes the internal stress field directly connected by the gravity of the fracture movement rock stratum and the area where the stress in the plastic failure area is less than the original stress; (ii) the high-stress area is the part where stress exceeds the original stress in the elastic–plastic area; (iii) the original stress area, that is, the area not affected by mining. The distribution characteristics of this area are shown in Figure 4. Under the condition of the same mining depths, the stress in the surrounding rock of the excavated roadway in different stress areas will directly determine the degree of damage to the rock surrounding the roadway. If the roadway is excavated in roadway 1, the stress level of the surrounding rock is lower, and under reasonable supporting conditions, the deformation of the surrounding rock can be effectively controlled. Serious highway deformation will unavoidably happen during excavation and maintenance if roadway 2 is excavated and is in a high-stress area. If the roadway is driven in the original rock stress zone, as shown in roadway 3, not only will a significant volume of coal resources be lost in the formation of the coal pillar, but when the mining depth reaches a certain value, the surrounding rock of the roadway will also be damaged [51,52,53].

2.4.2. Failure Process of Roadway Surrounding Rock

Before being affected by roadway excavation, coal and rock masses are in their original stress state. Subsequent excavation destroys the balance of the original stress, changing the stress environment around the roadway. Stress concentration occurs within a certain range of the roadway surrounding rock, with peak stress occurring near the coal wall. Coal enters the plastic failure state when the concentrated stress value is greater than the coal resistance limit. The stress peak will then shift to deep parts of the coal body and the pressure relief occurs within a certain range at the edge of the coal body. Failure in the coal wall initially occurs due to the failure of the two sides to deep areas of the coal side. With the continuous collapse and fragmentation of surrounding rock on the inner side of the roadway, the span of the roof and floor slab gradually increases, resulting in the roof and floor strata of the roadway forming transverse horizontal thrust under supporting pressure. Buckling failure will then occur due to horizontal reasoning. When the top and bottom plates are insufficient to resist the action of the transverse load, failure will occur [54,55,56]. This process is outlined in Figure 5.

3. Principle of Small Coal Pillar Roadway Protection and Surrounding Rock Control in Synergy

3.1. Deformation Failure Characteristics and Instability Mechanism of a Small Coal Pillar

3.1.1. Deformation and Failure Characteristics of a Small Coal Pillar

The deformation characteristics of the surrounding rock of gob-side roadways with small coal pillars can be described as:
  • Owing to the influence of secondary mining, mine ground pressure shows certain periodic regularity. During excavation, deformation along the gob side is greater than that along the solid-coal side, and during the mining of the roadway, deformation along the solid-coal side is greater than that along the gob side. For the gob-side roadway with a small coal pillar with moderately stable surrounding rock, the stress peak will be about 40 m from the front of the working face, having a leading influence distance of about 100 m. The surrounding rock of the roadway in the 0–10 m range in front of the working face will be severely deformed, and the relative displacement of the two sides will increase by about 10 times. The displacement of the roof and floor of the gob-side roadway increases 5–10 times more than that along the solid-coal roadway [57,58,59,60].
  • Due to differences in the depth an angle of dip of the coal seam, the compression direction and coal strength of the roadway differ. Failure characteristics and deformation differ as well. When the two sides of the roadway along the goaf are in the same working condition, the rock surrounding the roadway is subjected to asymmetrical loads, resulting in asymmetric deformation and failure [61,62,63].
  • The initial deformation rate is large. After roadway excavation, the initial deformation rate is large, taking place over a long time period under the action of mine pressure and having obvious rheological effects. If effective control measures are not taken, the deformation of the surrounding rock will continue to increase, eventually resulting in roadway instability and failure.

3.1.2. Instability Mechanism of a Small Coal Pillar

Gob-side entry advancement resulting in the redistribution of stope stress. Stress is mainly concentrated on the deep coal rock mass and the coal pillar. When stress in the surrounding rock exceeds its bearing limit, the surrounding rock mass will undergo plastic deformation and fail. Stope stress is redistributed by the stopping of the previous working face. When there are not enough overlying strata to withstand upper pressure and their own gravity, the strata will break and collapse onto the goaf. The fracture of the overlying strata will also cause pressure to transfer to the deep coal body. Stress formed under these conditions will be concentrated under the influence of lateral support pressure, and coal pillars adjacent to the goaf will be damaged due to excessive force. The roadway will therefore be excavated after reserving a certain width of coal pillar along the goaf side. During mining processes of a gob-side entry, the peak value of lateral bearing pressure continues to transfer to deep areas, increasing the influence range and thus resulting in stronger pressure. Under this influence, the coal pillar will produce plastic zone damage, and the damage on the coal pillar will seriously affect roadway stability. After understanding the failure mechanism of a coal pillar, we know that the instability failure of a coal body is mainly affected by strong pressure. When the coal body is subjected to small stress, the coal body has a certain bearing capacity, which will not change. With a gradual increase in load, cracks in the coal body will be compacted and their volume will be reduced; the compressive strength of coal and the rock mass will therefore be enhanced. Although the continued increase in applied pressure results in the plastic deformation of the coal and rock mass, the formed plastic block will still contain a certain bearing capacity. The plastic block will be completely destroyed when the pressure increase is greater than the maximum stiffness of the coal and rock mass. In relation to the coal pillar, because both sides of the coal pillar are free and there is no constraint control, the two sides of the unconstrained coal pillar are the first to be damaged when subjected to load. When mining is undertaken along the goaf, the coal pillar will undergo a stress concentration phenomenon. When the pressure exceeds the bearing limit of the coal pillar, the coal pillar will become unstable and collapse (Figure 6) [64,65].

3.2. Surrounding Rock Control in a Synergy Scheme and Parameter Design

In the light of the shortcomings of surrounding rock control, and based on the mechanism of roadway deformation and surrounding rock control ideas, surrounding rock control using a synergy scheme and parameters of coal pillar optimization (coal pillar size optimization)–roof cutting and pressure relief (roof presplitting blasting pressure relief)–grouting modification (coal pillar grouting reinforcement)–bolt support (high prestressed bolt cable support) in the 130,205 return air roadway in the Yangchangwan Coal Mine is proposed.

3.2.1. Coal Pillar Size Optimization Design

The formula to calculate the reasonable minimum width B of small coal pillars [66,67] is given as:
B = S 1 + S 2 + S 3
where S1 is the broken region produced in the small coal pillars along the goaf in this section.
S 1 = m A 2 tan φ 0 ln [ k γ H + C 0 / tan φ 0 tan φ 0 + P x / A ]
where m is the section excavation height, m; H is the buried depth of the roadway, m; γ is the average bulk density of overburden, MN/m3; K is the stress concentration coefficient; A is the lateral pressure coefficient; C0 is the cohesion of coal, MPa; φ0 is the coal internal friction angle, °; and Px is support resistance of coal seam, MPa.
S2 is the effective length of the upper anchor rod, m and the reinforcement coefficient is increased by 15%;
S3 is an increased coal pillar stability coefficient, m, calculated according to the value of (S1 + S2) 15%.
According to the geology of the extra-thick coal seam in the Yangchangwan Coal Mine, the parameter values are:
m = 3 m; H = 587.1 m; γ = 2.5×104 N/m3; K = 2.3; A = 0.33; C0 = 0.75 MPa; φ0 = 30°; Px = 0.
By using Equations (1) and (2), the reasonable coal pillar size was calculated to be 5.14–5.56 m. Combined with the section size of the roadway along the goaf (5.2 m), and the situation of fully mechanized caving, the width of the coal pillar on the return air roadway in 130,205 was designed to be 6 m and the roadway width was 5 m. On the basis of the author’s calculation, the coal pillar side of the roadway is completely within the range of the internal stress field (Figure 7).

3.2.2. Roof Blasting Pressure Relief Scheme Design

In order to reduce the weighting step behind the working face in 130,205, and to eliminate the hanging roof on the side of the 130,203 goaf, roof presplitting blasting was undertaken to cut the roof and unload pressure. An emulsion explosive with a ϕ50 × 500 mm powder barrel was used in blasting. An electric detonator was used in conjunction with a detonating cord to make the lead explosive. The drilling layout and charge quantity are described below.
The presplitting blasting layout parameters of the roof above the coal pillar are as follows: The advanced working face was 120 m long (the advanced distance was enlarged as much as possible). At 10 m intervals, holes were drilled with a diameter of 79 mm, and a depth of 20 m. Hole openings were positioned 1 m from the roadway. According to the site design scheme, boreholes were perpendicular to the direction of the roadway, and the elevation angle was 75° (Figure 8). The charge in each borehole was 22 kg, and the sealing length was not less than 7 m.

3.2.3. Small Coal Pillar Grouting Reinforcement

Grouting materials and additives used in this experiment included 425# ordinary silicate single-liquid cement slurry with a water–cement ratio of 0.7:1. This material had 8–10% ACZ-Ⅱ cement grouting added to it.
The layout of the grouting holes is as follows: Two rows of grouting holes were constructed on the upper side of the return airway. The upper row of holes was 0.7 m from the shoulder socket of the roof, and the lower row of holes was 1.5 m from the upper row of holes. The grouting holes’ diameter was 20–30 mm, depth was 2.5–3 m, and row spacing between the grouting holes was 1500 mm × 2000 mm. Cement and water glass were used for single-liquid grouting with a volume ratio of 1: 0.4. The holes were sealed using veil or cloth bags, and the grouting pressure was 1–1.5 MPa.
The grouting volume was calculated according to the grouting strength value method proposed [43] for ordinary cement grouting:
Qmax = λHRBnm
where Qmax is the maximum grouting volume per linear meter, m3; λ is the grouting loss coefficient; H is the roadway height, m; R is the grouting range, m; B is the filling coefficient; n is fissure rate; and m is the coal stone rate.
According to the specific working conditions of the 130,205 working face, λ = 1.2; H = 4 m; R = 3 m; B = 0.8; n = 15%; m = 1.
Using Equation (3), grouting volume per linear meter was about 1.728 m3. The length of the 130,205 grouting section was about 1100 m; thus, total grouting volume was 1900 m3.

3.2.4. Roadway Support Scheme Design

The roadway along the goaf was affected by multiple dynamic pressures, and the small coal pillar roadway was sealed by surface spraying to prevent weathering. Surface spraying also reduced strength attenuation speed, optimized support parameters, and ensured that the supporting capacity of the solid-coal seam and the small coal pillar on the roof were equivalent. The supporting capacity of the roof was equivalent to reduce the bias on the small coal pillars. The design scheme was as follows:
Combined with the site practice and the author’s calculation, a complete set of ϕ22 × 2500 mm rebar bolts were used on the roof of the roadway. Spacing between rows was reduced from 900 mm to 850 mm. A ϕ22 × 10,300 mm single steel strand anchor cable was adopted for the roof, and the anchor rod in the upper part of the roof was changed from the original ϕ20 × 2300 mm anchor rod to a complete set of threaded steel end anchor rods (ϕ22 × 2300 mm). Row spacing was set at 800 mm and the anchor part of the cable was ϕ22 × 6200 mm. A steel strand anchor cable (ϕ22 × 4300 mm) was also used along with top bolts round steel belts (processed from ϕ16 mm round steel), consistent with that used in the roadway section. With the side bolt, a W280-450-5 mm steel guard plate was used, and metal mesh (150 × 150 mm; 6000 × 1000 mm mesh size) was attached to the roof. High-strength plastic steel mesh (50 × 50 mm; 4000 × 1000 mm mesh size) was hung on the upper wall, and the mesh was interlocked together. A 70 mm thick layer of concrete was then sprayed onto the surface of the roadway (Figure 9).
A single bubble pressure relief ring was added to the bolt, and the pressure point and pressure-volume were set at 15 t and 25 mm, respectively. An anchor cable was added with a double bubble pressure relief ring, having a pressure point and pressure-volume of 20 t and 45 mm, respectively. The pallet adopted a bell-shaped pallet with a minimum thickness of 10 mm, and the anchor cable pallet adopted a 300 × 300 × 16 mm dish-shaped pallet. An alignment device was added to the pallet; the top plate of the diamond-shaped metal mesh and the two sides overlapped, and the overlap length was at least 100 mm; network spacing was at least 20 mm. The prestressing force of the top plate and the top anchor rod was at least 200 Nm, and the prestressing force of the anchor cable was at least 150 KN.

4. Investigating the Effect of Small Coal Pillars on Roadway Protection

4.1. Observation Scheme

4.1.1. Monitoring Scheme of Surrounding Rock Movement

The cross-point method was used in this study to monitor the displacement of the surrounding rock. Measuring nails were vertically installed in the middle of the roof and floor, and horizontally on the two sides of the 130,205 return air roadway. The height of the measurement points on two sides was adjusted according to the on-site situation; however, measuring points on the two sides had to be kept horizontal. According to the author’s calculation, the arrangement of observation points is shown in Figure 10a,b.
A total of three sets of stations were established and each station was set with three monitoring sections (each section was separated by 10 m): 1# station was situated 100 m from the open-off cut; 2# station was situated 150 m from the open-off cut; and 3# station was situated 300 m from the incision.

4.1.2. Monitoring of Axial Force Variation of the Anchor Cable

In order to monitor the variation of the anchor cable axial force, four groups of measuring stations were established (1#, 2#, 3#, and 4# measuring stations). The measuring stations were arranged in two stages: in the first stage, measuring stations 1# and 2# were arranged from the cut-out position of the working face. These measuring stations were established with three monitoring sections (each section was separated by 10 m). After the working face advanced to the small coal pillar stage, the second-stage measuring stations were arranged. Here, 3# and 4# stations were each equipped with one monitoring section (each section was separated by 50 m).
The 1# measuring station was situated 150 m from the open-off cut, being the influence range of the primary pressure on the main roof; the 2# station was arranged 300 m from the open-off cut, where the 130,205 working face connected with the 130,203 working face; the 3# station was situated 500 m from the open-off cut; the 4# station was situated 250 m from the entrance to the lane. On the basis of the author’s calculation, the equipment layout and on-site measuring points are highlighted in Figure 10 and Figure 11, respectively.

4.1.3. Monitoring Advanced Abutment Pressure

A ZYC-10 flexible hydraulic coal stress sensor and an infrared data acquisition instrument were adopted in this study. These instruments were used to monitor the variation law of the abutment pressure in the mining process of the working face. They were positioned in a refuge chamber more than 310 m from the open-off cut on the solid-coal side of the 130,205 return air roadway. The drilling holes for the stress-measuring points were numbered from the position of the open-off cut to the outside (1-1, 1-2, 1-3, for example). Borehole depth and diameter were 10 m and 48 mm, respectively. The stress sensor arrangement is shown in Figure 10a.

4.1.4. Detection of Fracture Morphology before and after Coal Pillar Grouting

A YS (B) mine-explosion-proof electronic drilling peeper was used to detect coal pillar failure before and after grouting. This instrument was also used to detect the gangues caving form of the 130,203 goaf. In the severely deformed area of the 130,205 return air roadway (130–230 m), coal body damage on one side of the coal pillar was detected using drilling holes. During normal mining practices on the working face, failure of the coal body on the solid-coal side and failure of the pillar coal are detected 50 m ahead of the working face under abutment pressure.

4.2. Implementation Effect

4.2.1. Mine Pressure Characteristics

After the coal pillar width in the 130,205 return air roadway was changed to 6 m, mine pressure behavior notably improved. According to field observation, the scenes of different areas of the roadway are shown in Figure 12.
Compared with a 35 m pillar under normal mine pressure conditions (shown in Figure 2), after the pillar width was changed to 6 m and the surrounding rock collaborative control scheme was implemented, the rock pressure recorded an obvious decrease. The deformation of the surrounding rock was effectively improved. Although sprayed concrete cracked under the action of local stress concentration, safety standards for the working face were met. Asymmetric deformation occurs in the roadway, and the deformation of the solid coal is large. However, the deformation of the roadway was in a controllable range, and deformation was significant 30 m ahead. As this deformation does not affect vehicles, pedestrians, or equipment, it, therefore, does not need to be rushed.

4.2.2. Convergence Deformation of the Roadway

According to the observation density and specified design requirements, continuous and systematic observations were carried out on the 130,205 return air roadway. The relationship curve between the cumulative approach and the advancing distance of the working face was calculated for observation sections 100 m, 150 m, and 300 m from the open-off cut of the working face.
According to on-site monitoring data, when the distance from station 1# to the working face was greater than 60 m, the deformation and deformation rate of the surrounding rock of the roadway was very small (Figure 13a). Distances between the two sides and the roof and floor were 20 mm and 17 mm, respectively, accounting for only 8.3% and 14.7% of total deformation. When the distance from the working face to the measuring station was less than 60 m, the deformation of the surrounding rock began to significantly increase, and the deformation rate increased. Cumulative deformation of the two sides of the roadway was 240.5 mm, and the cumulative deformation of the roof and floor was 115.3 mm. When the distance from station 2# to the working face was greater than 100 m, the deformation amount and deformation rate of the surrounding rock of the roadway were very small (Figure 13b). The displacement of two sides, and the roof and the floor, were 12 mm and 8.6 mm, respectively, accounting for only 5.3% and 4.6% of the total deformation amount. When the distance from the working face to the measuring station was less than 100 m, the deformation of the surrounding rock began to significantly increase; the accumulated deformation of the two sides of the roadway was 225.5 mm, and the accumulated deformation of the roof and floor was 186.7 mm. Station 3# slowly deformed 130 m from the working face, and mining at the working face resulted in a gradual increase in the deformation of the surrounding rock (Figure 13c). The cumulative deformation of two sides of the roadway was 193.9 mm, and the cumulative deformation of the roof and floor was 93.4 mm.
Results from each station indicate that the displacement of both sides of the roadway was greater than that of the roof and floor. The deformation of the roadway gradually increased with the mining of the working face (Figure 13). Mining mostly caused the surrounding rock of the roadway to deform; specifically, within 60 m of the working face, the surrounding rock of the roadway started to visibly deform, and the approaching speed peaked at 20 m. This indicates that the advanced influence range in the mining process of the working face was roughly within 100 m, and the distance of the roof deformation significantly increased by 25 m in front of the work; however, it sharply increased around 18 m.

4.2.3. Axial Force Variation of the Anchor Cable

According to our design requirements, the axial force of the bolt (cable) in the 130,205 return air roadway was continuously and systematically observed. Due to space limitations, data from three observation sections (150 m, 160 m, and 300 m) from the open-off cut of the working face were analyzed.
According to on-site monitoring data, the relationship between the variation of the axial force of the bolt (cable) and the advancing distance of the working face (Figure 14) indicated that bolt (cable) stress began to significantly increase at 30–40 m ahead of the working face. As the working face advanced, the increase in bolt (cable) stress had an increasing trend, with the last 10–15 m bolt (cable) stress growing the fastest. This phenomenon has an important relationship with the distribution of abutment pressure in front of the working face. The roadway surrounding the rock moved closer in advance to increase the stress of the anchor bolt (cable), preventing the deformation of the coal and rock surface. In the last 0–10 m, the stress of some bolts (cable) slightly decreased. In this range, coal and rock mass were crushed and destroyed under strong abutment pressure. The development of internal cracks increased, and the pulling and hoisting efficiency of the bolt cable slightly decreased. Due to the large tray and anchor net with a large protective surface area, the two sides moved within a controllable range, and there was no large area of coal spray or falling material on the side. Few pull-out failure phenomena of the anchor bolt were recorded, and the pull and hanging performance of anchor bolts were always effective, ensuring good control of the rock surrounding the roadway.

4.2.4. Distribution of Advanced Abutment Pressure

In accordance with the design requirements for abutment pressure monitoring in front of a working face, continuous systematic observations were undertaken for the 130,205 return air roadway. According to on-site monitoring data, the relationship between abutment pressure change in front of the working face and the advanced distance of the working face was monitored using borehole stress gauges installed 310 m, 312 m, and 316 m from the opening of the working face (Figure 15).
The results indicated that stress increased closer to the working face, away from the station, and, apart from 60–80 m in front of the working face, abutment pressure gradually increased. After the peak was reduced, an irregular parabola-shaped face abutment pressure curve was recorded. The peak on the left side of the display reflected an elastic compression state, and the peak on the right coincided with a plastic compression state of the coal. The stress concentration coefficient (K) was 2.5–3. The plastic zone was within 20–30 m, and coal in this area was crushed and destroyed. This coal served as a protective belt for the impact ejection of coal at the working face, reducing the possibility of impact damage. The coal body was in an elastic stress state within 30–60 m in front of the working face, and large compression elastic energy was present in the coal body, having a large compression effect on the floor, possibly leading to the buckling failure of the floor rock mass.

4.2.5. Detection of Fracture Morphology in the Coal Pillar

According to field observation, fracture development in the coal pillar before and after grouting (Figure 16) indicated that, before the coal pillar was grouted, the whole wall was rough and the degree of fragmentation was relatively high (Figure 16a). The top of the borehole was visibly penetrated by a large crack, and the integrity of the coal body was poor. After grouting (Figure 16b), original cracks were filled with cement slurry, the walls were smooth, and integrity was better. By comparison, it can be seen that cracks in the coal body can be better filled using the grouting method. Here, the integrity of the coal body increased and its bearing capacity increased.

5. Discussion

The results from this analysis indicate that the large deformation of the large-section dynamic pressure roadway in the extra-thick coal seam was due to interactions between the stress environment, the mechanical strength of the surrounding rock, and the support method. From these three aspects, the control method affecting the determination of the surrounding rock was considered.
According to deformation characteristics of rock surrounding a roadway along a goaf with small coal pillars, the rock of the surrounding roadway was mainly controlled by the following means: Firstly, by optimizing the size of the coal pillar, the roadway was arranged in the internal stress field with a low stress level to improve the stress environment of the roadway. Secondly, roof presplitting blasting can eliminate the side hanging roof of the goaf and reduce the potential risk of large area caving, as well as reducing the risk of roadway impact. Thirdly, the discontinuous structural plane inside the coal body was filled by grouting, and rock masses on both sides of the structural plane were “bonded” together to increase the strength of the structural plane, thus improving the integrity of the rock mass and enhancing its resistance. Fourthly, the stability of the surrounding rock was maintained through high prestress, high strength, and high elongation of the anchor cable, thereby controlling the separation of coal strata, sliding, and other discontinuous, uncoordinated expansion deformation.
The parameters used to control the surrounding rock in the synergy scheme were designed in accordance with geological conditions present in the Yangchangwan Coal Mine. Firstly, using the limit equilibrium theory and roadway section size, the reasonable width of a small coal pillar was determined to be 6 m. The drilling layout, charging amount, coal pillar grouting amount, and bolt cable support parameters of the roof blasting pressure relief hole were also determined. After implementing surrounding rock control in the synergy scheme, the deformation of the two sides was still recorded to be greater than that of the top and bottom plates. The maximum subsidence of the top plate was 55 mm, the maximum bottom heave of the bottom plate was 55 mm, and the maximum values of the upper and lower side drums were 30 mm and 70 mm, respectively (Figure 13). It was seen that the upper side drum of the air roadway recorded the smallest change, indicating that a 6 m small coal pillar can effectively maintain roadway stability. By comparing the results in Figure 3 and Figure 10, it can be seen that, compared with the condition of the 35 m coal pillar, the breaking rate of the anchor cable under the condition of a 6 m coal pillar was reduced by more than 90%, and the deformation of the rock surrounding the roadway was reduced by more than 70%.
In conclusion, after implementing the surrounding rock control in the synergy of coal pillar optimization (coal pillar size optimization)–roof cutting and pressure relief (roof presplitting blasting pressure relief)–grouting modification (coal pillar grouting reinforcement)–bolt support, the surrounding rock deformation of the mining roadway was effectively controlled to meet the requirements of underground production.

6. Conclusions

The deformation control of rock surrounding a fully enlarged cross-section mining roadway in an extra-thick coal seam of a western coal mine was examined. Using the 130,205 mining roadway in Yangchangwan Coal Mine as a case study, factors affecting the deformation of the surrounding rock were explored using theoretical analysis and industrial experiments. A “coal pillar optimization-roof cutting destressing-grouting modification-rock bolting” surrounding rock control method in synergy was explored. The main conclusions from this analysis were as follows:
  • The large deformation of the rock surrounding a large section of mining roadway occurred due to the interaction of high stress, surrounding rock strength, and the support mode. The deformation of the rock surrounding the roadway presents obvious asymmetry and a strong rheological effect. Strong floor heave and two side movements are the main ground pressure characteristics of this kind of mining roadway.
  • The original bolt mesh shotcrete support cannot effectively control the large deformation of the surrounding rock of a fully mechanized top-coal caving face in an extra-thick coal seam. The support body damage was mainly manifested as the breakage of the bolt cable and failure under the action of composite stress; sprayed concrete cracks and spalls under the action of local concentrated stress; and the I-beam, etc., was bent and deformed under the action of concentrated stress.
  • The results from this study enable a suitable surrounding rock control in synergy scheme to be proposed, as well as parameters for coal pillar optimization–roof cutting destressing–grouting modification–rock bolting. Field experiments and observational information on surrounding rock deformation support the implementation of stress and grouting effects undertaken. The monitoring results indicate that, compared with the original surrounding rock control plan, the implementation of the surrounding rock control in a synergy scheme resulted in the anchoring effect of the anchor rod cable being fully exerted. The breaking rate was reduced by more than 90% and the deformation of the rock surrounding the roadway was reduced by about 70%. Field practice of the roadway surrounding rock control in the synergy method indicated that the rock deformation was effectively controlled, and the successful application of this technology was able to provide reliable technical and theoretical support for the Ningdong mining area and mines with similar conditions. These results have good promotional and application value.

Author Contributions

Methodology, H.B. and G.Z.; Formal analysis, A.C.; Investigation, L.P.; Resources, Z.W.; Data curation, J.H.; Project administration, J.T.; Funding acquisition, X.L. All authors have read and agreed to the published version of the manuscript.

Funding

The study was financially supported by the Taishan Scholars Project, National Natural Science Foundation of China (Grant Number. 52174121 and 52104204), Natural Science Foundation of Shandong Province (ZR2021QE170), Project of Shandong Province Higher Educational "Youth Innovation Science and Technology Plan" Team (Grant Nos. 2021KJ060).

Data Availability Statement

The data presented in this study are available on request from the corresponding author. The data are not publicly available due to the confidentiality of the project involved.

Acknowledgments

We thank the anonymous reviewers for their constructive feedback.

Conflicts of Interest

The authors declare no conflict of interest.

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Figure 1. Mining location and side-view layout of the study site. (a) Mine location; (b) Side-view layout of the study site.
Figure 1. Mining location and side-view layout of the study site. (a) Mine location; (b) Side-view layout of the study site.
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Figure 2. Original bolt cable support scheme of return air roadway.
Figure 2. Original bolt cable support scheme of return air roadway.
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Figure 4. Stope abutment pressure distribution.
Figure 4. Stope abutment pressure distribution.
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Figure 5. Instability failure process of the roadway. (a) Initial form before destruction; (b) Initial form after destruction.
Figure 5. Instability failure process of the roadway. (a) Initial form before destruction; (b) Initial form after destruction.
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Figure 6. Small coal pillar instability mechanism.
Figure 6. Small coal pillar instability mechanism.
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Figure 7. The original design of the mining roadway position.
Figure 7. The original design of the mining roadway position.
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Figure 8. Diagram of blasting arrangement for side borehole of coal pillar. (a) Lateral view; (b) Plan view.
Figure 8. Diagram of blasting arrangement for side borehole of coal pillar. (a) Lateral view; (b) Plan view.
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Figure 9. Design optimization of support for 130,205 return air roadway.
Figure 9. Design optimization of support for 130,205 return air roadway.
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Figure 10. Arrangement of observation instruments. (a) Plan view; (b) Section view.
Figure 10. Arrangement of observation instruments. (a) Plan view; (b) Section view.
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Figure 11. Field of axial force monitoring of bolt (cable).
Figure 11. Field of axial force monitoring of bolt (cable).
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Figure 12. Mine pressure behavior. (a) 15 m ahead of working face; (b) 10 m ahead of working face; (c) 30 m ahead of working face; (d) 30 m ahead of working face.
Figure 12. Mine pressure behavior. (a) 15 m ahead of working face; (b) 10 m ahead of working face; (c) 30 m ahead of working face; (d) 30 m ahead of working face.
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Figure 13. Deformation of surrounding rock of roadway. (a) 1# roadway surface displacement observation curve; (b) 2# roadway surface displacement observation curve; (c) 3# roadway surface displacement observation curve.
Figure 13. Deformation of surrounding rock of roadway. (a) 1# roadway surface displacement observation curve; (b) 2# roadway surface displacement observation curve; (c) 3# roadway surface displacement observation curve.
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Figure 14. Force curve of anchor cable in 130,205 return airway. (a) Bolt stress at 150 m and 160 m from open cut; (b) Bolt stress at 300 m from the open cut.
Figure 14. Force curve of anchor cable in 130,205 return airway. (a) Bolt stress at 150 m and 160 m from open cut; (b) Bolt stress at 300 m from the open cut.
Energies 15 09040 g014aEnergies 15 09040 g014b
Energies 15 09040 g015 550
Figure 15. Force curve of 130,205 return airway borehole stress gauge.
Figure 15. Force curve of 130,205 return airway borehole stress gauge.
Energies 15 09040 g015
Energies 15 09040 g016 550
Figure 16. Failure characteristics of coal pillars (a) before and (b) after grouting.
Figure 16. Failure characteristics of coal pillars (a) before and (b) after grouting.
Energies 15 09040 g016
Table
Table 1. Coal strata distribution and geological description.
Table 1. Coal strata distribution and geological description.
LithologyThickness (m)RemarksLithology Description
Siltstone10.4Main floorGriseous, silty structure, block structure
Coarse sandstone39.27Overlying strataGray-white, medium-particle structure, block structure, poor sorting
1# coal seam0.92Overlying strataBlack, lump, dark briquette, Eye-shaped fracture
Fine sandstone4.96Main roofBlack-gray, mainly composed of quartz
Medium sandstone12.7Main roofGray-white, mainly composed of feldspar minerals and mud debris
Fine sandstone3.23Immediate roofGray, dark gray, mainly composed of feldspar minerals and mud debris
Siltstone4.62Immediate roofCharcoal gray, upper intercalated with thin sandstone
Carbon mudstone0.4False roofBlack, containing a small amount of siltstone and pyrite
2# coal seam9Coal seamBlack, lumpy
Table
Table 2. Mechanical parameters of coal and rock mass.
Table 2. Mechanical parameters of coal and rock mass.
LithologyTensile Strength/MPaElastic Modulus/MPaPoisson Ratio/MPaCohesion/MPaAngle of Internal Friction/°
Siltstone4.5214.60.24.6632
Medium sandstone3.8614.40.234.8326
Fine sandstone5.17.520.2251.728
Siltstone3.5712.50.24.2730
Carbon mudston1.710.750.21123.8
2# coal seam1.31.860.260.7530
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Hao, J.; Chen, A.; Li, X.; Bian, H.; Zhou, G.; Wu, Z.; Peng, L.; Tang, J. Analysis of Surrounding Rock Control Technology and Its Application on a Dynamic Pressure Roadway in a Thick Coal Seam. Energies 2022, 15, 9040. https://doi.org/10.3390/en15239040

AMA Style

Hao J, Chen A, Li X, Bian H, Zhou G, Wu Z, Peng L, Tang J. Analysis of Surrounding Rock Control Technology and Its Application on a Dynamic Pressure Roadway in a Thick Coal Seam. Energies. 2022; 15(23):9040. https://doi.org/10.3390/en15239040

Chicago/Turabian Style

Hao, Jian, Anfa Chen, Xuelong Li, Hua Bian, Guanghua Zhou, Zhenguo Wu, Linjun Peng, and Jianquan Tang. 2022. "Analysis of Surrounding Rock Control Technology and Its Application on a Dynamic Pressure Roadway in a Thick Coal Seam" Energies 15, no. 23: 9040. https://doi.org/10.3390/en15239040

APA Style

Hao, J., Chen, A., Li, X., Bian, H., Zhou, G., Wu, Z., Peng, L., & Tang, J. (2022). Analysis of Surrounding Rock Control Technology and Its Application on a Dynamic Pressure Roadway in a Thick Coal Seam. Energies, 15(23), 9040. https://doi.org/10.3390/en15239040

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