Roadway Surrounding Rock under Multi-Coal-Seam Mining: Deviatoric Stress Evolution and Control Technology

Multi-coal-seam mining creates surrounding rock control diﬃculties, because the mining of a coal face in one seam can aﬀect coal faces in another. We examine the eﬀects of multi-coal-seam mining on the evolution of the deviatoric stress distribution and plastic zone in the roadway surrounding rock. In particular, we use numerical simulation, theoretical calculation, drilling detection, and mine pressure observation to study the distribution and evolution characteristics of deviatoric stress on Tailgate 8709 in No. 11 coal seam in Jinhuagong mine when the N8707 and N8709 coal faces in No. 7-4 coal seam and the N8707 and N8709 coal faces in No. 11 coal seam are mined. The evolution laws of deviatoric stress and the plastic zone of roadway surrounding rock in the advance and behind sections of the coal face are studied, and a corresponding control technology is proposed. The results show that the peak value of deviatoric stress increases with the advance of the coal face, and the positions of the peak value of deviatoric stress and the plastic zone become deeper. The deﬂection angle of the peak stress after mining at each coal face and the characteristics of the peak zone of deviatoric stress and the plastic zone of the roadway surrounding rock under the disturbance of multi-coal-seam mining are determined. In conclusion, the damage range in the roadway roof in the solid-coal side and coal pillar is large and must be controlled. A combined support technology based on high-strength and high pretension anchor cables and truss anchor cables is proposed; long anchor cables are used to strengthen the support of the roadway roof in the solid-coal side and coal pillar. The accuracy of the calculated plastic zone range and the reliability of the combined support technology are veriﬁed through drilling detection and mine pressure observation on site. This research can provide a point of reference for roadway surrounding rock control under similar conditions.


Introduction
Multi-coal-seam mining is common in China [1,2], where coal seams that are shallow or have good mining conditions are nearly depleted in a short service life with the continuous increase in coal mining intensity [3]. During the downward mining of multiple coal seams, the supporting pressure above the residual coal pillars in the upper section and the collapsed roof in the gob will cause damage to the floor, which is now affected not only by the original rock stress but also by mining stress. Over time, the concentration, transmission, and change of stress in the floor due to multi-coalseam mining will affect the panel of the lower coal seam, and it will become significantly more difficult to control the surrounding rock of the roadway [4,5]. Much valuable research has been conducted on developing mechanisms and technology for the control of surrounding rock influenced by multi-coal-seam mining in recent years. For example, Zhu et al. [6] proposed a classification criterion for gob caving under multi-coal-seam mining conditions based in part on the specifics of the interior load-bearing structures. Tan et al. [7] studied the failure characteristics of the upper strata after multi-coal-seam mining and showed that when the mining height of the lower coal seam is large, the strata of the upper coal seam have many fractures. Li et al. [8] detected the development of cracks in the floor of the upper coal seam after the mining of the lower coal seam. Qiang et al. [9] used a geographic information system to evaluate the likelihood of water inrushes and showed that different coal seams have different vulnerabilities in multiseam mining conditions. Zhang et al. [10] addressed the peak stress and influencezone location for concurrent pillar recovery in two coal seams with 21 m of interburden at an approximate depth of 305 m. Deutsch and Wilde [11] modeled multiple coal seams in three dimensions without surface-contour or isochore mapping. Shang et al. [12] investigated the chain pillar and gateroad failure process, in which there is an increase in stress in the lower coal seam caused by a residual coal pillar in the upper coal seam.
According to the theory of plastic mechanics, the surrounding-rock stress is the superposition of spherical stress and deviatoric stress. e deviatoric stress is the root cause of the distortion of the roadway surrounding rock, and it plays a major role in the occurrence and development of plastic deformation [13,14]. ere have been several observational and theoretical studies of the distribution of the deviatoric stress and the plastic zone of roadway surrounding rock. For example, Ma et al. [15] applied the Mohr-Coulomb criterion to obtain the deviatoric stress formula for circular-roadway surrounding rock in a nonuniform stress field. Xu et al. [16,17] simulated the deviatoric stress distribution in ultrahigh roadway and under abandoned coal pillars in multicoal-seam mines. Yu et al. [18] deduced the analytical form of the deviatoric stress field of surrounding rock from rock mechanics and other related theories and analyzed the distribution of surrounding rock under different conditions. He et al. [19] investigated the roof deviatoric stress, surrounding-rock fracture field, and deformation law when changing the width of a large-section open-off cut roof from 6 to 10 m. Xie et al. [20,21] adopted a strain-softening model to simulate the deviatoric stress evolution laws of surrounding rock at gob-side entry retaining throughout the whole process of the coal face's advance.
Most existing research on roadway surrounding-rock failure in multi-coal-seam mining has focused on the vertical or horizontal stresses. ere has been no targeted study on the distribution law of deviatoric stress in roadway-floor surrounding rock disturbed by multi-coal-seam mining.
is paper considers the effect of multi-coal-seam mining on Tailgate 8709 in Jinhuagong mine, Datong City, Shanxi Province, China. e FLAC 3D numerical calculation software is used to study the disturbance mechanisms, deviatoric stress, and plastic zones of the N8707 coal face and the N8709 coal face in No. 7-4 coal seam and the N8707 coal face and the N8709 coal face in No. 11 coal seam. e peak deviatoric stress profile and plastic zone profile of the roadway surrounding rock are determined, and a combined support technology using high-strength and high pretension anchor cables and truss anchor cables is proposed for Tailgate 8709. e accuracy of the plastic zone and the reliability of the combined support technology are verified by arranging on-site drilling detection and mine pressure observation. is research can provide a point of reference for the control of surrounding rock similarly disturbed by multi-coal-seam mining.

Project Overview
Jinhuagong coal mine is a short distance west of Datong City, Shanxi Province, China. It is a typical multi-coal-seam mine with complex geological conditions. e thicknesses of No. No. 11,and No. 12

Numerical Simulation of Tailgate 8709 under
Multi-Coal-Seam Mining

Numerical Model.
e FLAC 3D numerical simulation software has been used to construct a mathematical model of Tailgate 8709 in No. 11 coal seam in Jinhuagong mine under multi-coal-seam mining. e model is visualized in Figure 2. We simulate each rock layer using the Mohr-Coulomb constitutive model, the one most widely used in FLAC 3D numerical simulation of various rock formations. Following [22][23][24][25][26][27][28][29][30][31], we use a strain-softening model to simulate the mechanical behavior of coal seams. Like [32][33][34][35][36][37][38][39], we use the natural collapse method to deal with the gob of the coal face in the simulation. e coal face in the model is taken to extend 200 m along the x-axis. e y-axis represents the advancing direction of the coal face, taken to extend 120 m, and the z-axis is the vertical direction, taken to extend 140 m.
e size of the model is thus 200 m × 120 m × 140 m. Because various joints and cracks often exist in the rock mass on-site, the rock mechanical parameters measured in laboratory tests are significantly higher than the actual parameters in the field; laboratory results cannot be used directly as the mechanical parameters of each rock layer in the model. Based on the Hoek-Brown failure criterion [40,41], the laboratory test results are calculated and processed to obtain the mechanical parameters of various rock masses listed in Table 1.
e mining process of the simulated coal face is required to be consistent with the actual mining sequence at Jinhuagong mine; that is, the N8707 and N8709 coal faces in No. 7-4 coal seam and the N8707 and N8709 coal faces in No. 11 coal seam are mined in the stated order. In the simulation, the excavation step is set as 5 m. Nephograms of the deviatoric stress in the roadway surrounding rock at 2 Advances in Civil Engineering different stages are analyzed, and deviatoric stress data is extracted to draw stress curves. e disturbances to Tailgate 8709 in No. 11 coal seam caused by multi-coal-seam mining are explored with nephograms of deviatoric stress and models of the plastic zone distribution.

e Strain-Softening Model and Its
Parameters. e plastic shear strain parameter ε ps of coal after yielding used in the strain-softening criterion is where ε p 1 and ε p 3 are the plastic principal strain components, and e strain-softening model employs the Mohr-Coulomb yield criterion [42]:

Advances in Civil Engineering
where σ 1 and σ 3 are the maximum and minimum principal stresses, respectively; φ(ε ps ) is the friction angle after plastic failure: and c(ε ps ) is the cohesion after plastic failure: ε r is the plastic strain in the residual stage; ε p is the plastic strain at the peak stress; φ p is the friction angle at the peak stress; c p is the cohesion at the peak stress; φ r is the friction angle in the residual stage; c r is the cohesion in the residual stage.
In order to verify the rationality of adopting a strainsoftening model for the coal seam in the FLAC 3D software, a standard cylinder test piece, Φ50 × 100 mm, was fabricated, and a load was uniformly applied to its upper part at a constant rate until the test piece was damaged. is experiment was then simulated using FLAC 3D . Figure 3 shows that the stress-strain curves from laboratory experiment and numerical simulation are consistent, indicating that it is reasonable to use a strain-softening criterion in the model.

Distribution and Evolution Characteristics of Deviatoric Stress in Roadway Surrounding Rock Affected by Multi-Coal-Seam
Mining. In order to study the disturbance of the N8707 and N8709 coal faces in No. 7-4 coal seam on the N8707 and N8709 coal faces in No. 11 coal seam on Tailgate 8709, the distribution and evolution laws of deviatoric stress and the plastic zone of the roadway surrounding rock in the mining process of each coal face are studied.

Overview of Deviatoric Stress.
When the surrounding rock in a coal mine is in a state of hydrostatic pressure, the principal stresses σ i (i � 1, 2, 3) are perpendicular to each    other and equal in value: σ 1 � σ 2 � σ 3 . e rock stress in the roadway will change after the roadway is excavated, and the following relationship applies: σ 1 ≥ σ 2 ≥ σ 3 . It can also be shown that where σ m is the average stress: σ m � (1/3) 3 i�1 σ i . e deformation of an object is caused by changes in its volume and shape. e second term on the left side of formula (6) is the spherical stress tensor, which causes changes in volume, while the first item is the deviatoric stress tensor, which causes changes in shape (plastic deformation and failure). e deviatoric stress is the essential cause of rock-mass failure. e principal deviatoric stress defined as σ i − σ m in the above formula is principal deviatoric stress. e maximum principal deviatoric stress (σ ′ ) is therefore given by e shear stress of the shallow surrounding rock in the roadway is concentrated after the roadway is excavated, and the surrounding rock releases energy, unloading deformation. e deviatoric stress is a combination of horizontal, vertical, and tangential stress that can be used to characterize the distribution of shear stress, the essential source of rock deformation and failure, under the action of a load. e present research therefore highlights deviatoric stress as an index reflecting the failure evolution of surrounding rock more comprehensively than does vertical or horizontal stress alone. Figure 4 shows the deviatoric stress distributions after the N8707 and N8709 coal faces in No. 7-4 coal seam have been mined. It can be seen from Figure 4(a) that an approximately "fan shaped" peak zone of deviatoric stress is formed at the right coal-pillar boundary after the N8707 coal face in No. 7-4 coal seam is mined out and that the deviatoric stress gradually decreases away from the peak zone. e surrounding rock of Tailgate 8709 in No. 11 coal seam also has a peak zone of deviatoric stress (concentrated mainly in the roof and floor), and the stress of the roadway surrounding rock has an approximately symmetrical distribution; this shows that the roadway is not obviously affected by the mining disturbance of the N8707 coal face in No. 7-4 coal seam. Because Tailgate 8709 in No. 11 coal seam is 38.8 m below the N8709 coal face in No. 7-4 coal seam, it is far away from the N8707 coal face in No. 7-4 coal seam and is therefore not significantly disturbed by the mining of the N8707 coal face in the upper coal seam. Figure 4(b) shows that an approximately symmetrical "八" shaped peak zone of deviatoric stress, with the left and right sides of coal pillars as the boundary, is formed after both sides of the coal pillar in No. 7-4 coal seam are mined. e stress value below the peak zone of deviatoric stress falls off slowly. e peak zone of deviatoric stress of Tailgate 8709 in No. 11 coal seam deflects obviously compared with Figure 4(a). e reason for this phenomenon is that Tailgate 8709 is located under the N8709 coal face in No. 7-4 coal seam, and the partial stresses of two overlap. e peak zone of deviatoric stress of the roadway surrounding rock deflects to the bottom corner on the coal-pillar side and the shoulder corner of the roadway roof on the solid-coal side. Figure 5 shows that the distribution curves of deviatoric stress in the surrounding rock in solid coal, coal pillar, roof, and floor are approximately the same after the N8707 and N8709 coal faces in No. 7-4 coal seam are mined. e deviatoric stress of the roadway surrounding rock, especially of the roof, shows an increasing trend after the N8709 coal face in No. 7-4 coal seam is mined; the peak value of deviatoric stress of the roof (3.73 MPa after mining of the N8707 coal face) increases to 4.06 MPa after mining of the N8709 coal face. e farther one looks away from the roof of Tailgate 8709 in No. 11 coal seam, the more obvious is the increase of deviatoric stress. erefore, it can be concluded that the deviatoric stress in surrounding rock of Tailgate 8709 in No. 11 coal seam is redistributed due to the fact that the N8709 coal face in No. 7-4 coal seam is mined combining with the deflection of the peak zone of deviatoric stress in roadway surrounding rock in Figure 4(b), which makes deviatoric stress of roadway surrounding rock gradually increase. After the N8709 coal face in No. 7-4 coal seam is mined, the peak distances of deviatoric stress in solid-coal side, coal pillar, and floor of Tailgate 8709 in No. 11 coal seam are about 2.0 m, 2.5 m, and 2.0 m, respectively. e peak distance of the roof after mining of the N8707 coal face increases from 2.5 m to 4.0 m after mining of the N8709 coal face. e roadway roof is the area most affected by mining of the coal face in the overlying coal seam.

3.3.3.
e Deviatoric Stress Distribution of Tailgate 8709 under the Mining Disturbance of the N8707 Coal Face in No. 11 Coal Seam. Figure 6 shows nephograms of the deviatoric stress distribution of surrounding rock in the advance and behind sections of Tailgate 8709 when the N8707 coal face in No. 11 coal seam is mined for various distances. e nephograms and curves with positive numbers represent the advance sections and those with negative numbers the behind sections. It can be seen that the nephograms of deviatoric stress of Tailgate 8709 are approximately the same, which shows the rationality of leaving 25 m coal pillars between the N8707 coal face and the N8709 coal face in No. 11 coal seam. With a sufficient width of coal pillar, mining of the adjacent coal face has no great impact on Tailgate 8709. e peak zone of deviatoric stress is mainly at the shoulder corner of the roadway roof in solid coal and at the bottom corner of the coal pillar.
e peak stress at the shoulder corner of the roadway roof in solid coal is far from the roadway surface, so it is necessary to strengthen the support for it in solid coal.   e main conclusion to be drawn is that the farther the behind coal face is, the greater the deviatoric stress of Tailgate 8709 is.
at is, the deviatoric stress of roadway surrounding rock is greater than that of roadway surrounding rock at different positions in the advance section when the behind coal face is at a distance of 40 m. e peak values of deviatoric stress in the roof, solid coal, coal pillar, and floor of Tailgate 8709 at 40 m behind the coal face are about 4.11 MPa, 7.17 MPa, 5.60 MPa, and 4.75 MPa, respectively, and the distances between the peak stress locations and the roadway surface are about 4.0 m, 2.5 m, 4.0 m, and 2.5 m. In the advancing coal-face section, the deviatoric stress of Tailgate 8709 shows little change, which indicates that the mining of the N8707 coal face in No. 11 coal seam has little effect on deviatoric stress in the surrounding rock of Tailgate 8709 in the advance section, although it causes some disturbance in the deviatoric stress distribution of roadway surrounding rock in the behind coal-face section. e peak positions of deviatoric stress in solid coal, coal pillar, roof, and floor of roadway in the advance and behind sections exhibit no obvious changes. erefore, Tailgate 8709 is hardly affected by the mining of the N8707 coal face in No. 11 coal seam, proving that it is reasonable to set a 25 m coal pillar between two adjacent coal faces. e deviatoric stresses at 2 m (in the shallow surrounding rock of the solid-coal side), 3 m (in the shallow surrounding rock of the coal pillar), and 4 m (in the shallow surrounding rock of the roof ) show an increasing trend, which indicates that the shallow surrounding rock of the roadway is damaged to a certain extent and the deviatoric stress is low. us, it is necessary to control the stability of rock masses with fracture development in this range, while at the same time strengthening the support for the shoulder corner of the roadway roof on the solid-coal side.   Figure 8 shows nephograms of deviatoric stress at different positions of the advanced region of Tailgate 8709 when the N8709 coal face in No. 11 coal seam is mined to various distances. It can be seen that the deviatoric stress of the surrounding rock of Tailgate 8709 increases rapidly when approaching the coal face gradually, and the roadway surrounding rock has an irregularly distributed peak zone of deviatoric stress. e position of the peak zone of deviatoric stress has been deflected and deepened compared with that in Figure 6, and the peak zone is stably distributed at the shoulder corner of the roadway roof on the solid-coal side and the bottom corner of the coal pillar.
According to Figure 9, the deviatoric stress curves of Tailgate 8709 in solid-coal side, coal pillar, roof, and floor differ at different positions of the advance coal face. Deviatoric stress values on the solid-coal side near the coal face gradually increase with depth. e deviatoric stress curves in this region rise at first, when the advance coal face is far away, then decrease, and finally tend to become flat. e deviatoric stress curves in the coal pillar show an "M" shape. e farther the advance coal face is, the less the deviatoric stress is. e deviatoric stress curves on the solid-coal side and in the coal pillar will gradually become coincident after 40 m advance of the coal face. is shows that, over this much advance of the coal face, the mining influence is gradually reduced, and the distribution of deviatoric stress tends to become stable. e deviatoric stresses of the roof and floor near the coal face are greatly affected by the mining of the coal face. e distance between the peak location of deviatoric stress of the roof and surface of the roof increases to 4.0 m, and the position of the  Tailgate  8709 under Multi-Coal-Seam Mining. Figure 10 shows nephograms of the deviatoric stress and the plastic zone of the surrounding rock in Tailgate 8709 after the mining of No. 7-4 and No. 11 coal seams. From the nephograms (Figure 10(a)), it can be concluded that the peak zone of deviatoric stress of the roof of the N8709 coal face in No. 7-4 coal seam is deflected by 34°from the peak zone of deviatoric stress of the surrounding rock in Tailgate 8709. e peak zone of deviatoric stress of the roof continues to be deflected by 13°after mining of the N8707 coal face in No. 11 coal seam, and the peak zone of deviatoric stress of the roof continues to be deflected by 53°when No. 11 coal seam is 10 m ahead of the N8709 coal face. erefore, the peak zone of deviatoric stress of Tailgate 8709 continually deflects to find a new effective bearing structure after the roadway surrounding rock is damaged under the influence of multicoal-seam mining. It should be ensured during the support design that the anchor cable can pass through the position of the peak zone of deviatoric stress (Figure 10(d)). Figure 10(a) shows that the plastic zone of the surrounding rock of Tailgate 8709 is symmetrically distributed after the mining of the N8707 coal face in No. 7-4 coal seam, which shows that Tailgate 8709 in No. 11 coal seam is scarcely affected by the mining of the N8707 coal face in No. 7-4 coal seam.

Distribution and Evolution Laws of Plastic Zone in
is is approximately consistent with the Under a nonuniform stress field, when the lateral pressure coefficient is not equal to 1, the relationship between the radius r of the plastic zone of the roadway and the deviatoric stress difference can be calculated by the following formula [15]: where a is the equivalent radius of the roadway, taking the outer circle radius of the roadway as 2.65 m; c is the bulk density of the rock, 25 kN/m 3 ; H is the buried depth of the roadway, 300 m; φ is the internal friction angle, 27°; λ is the coefficient of lateral pressure, 1.2; C is the cohesion, 0.9 MPa; s 1 is the maximum principal deviatoric stress, 6.98 MPa; s 3 is the minimum principal deviatoric stress, 1.51 MPa; θ is the angle of any point in the plane. is formula shows that the maximum radius of the plastic zone of the circular roadway is 6.45 m at 10 m ahead of the N8709 coal face in No. 11 coal seam. e depth of the plastic zone of the roadway is 3.80 m after subtracting 2.65 m from the radius of circular roadway. e result from the formula is approximately consistent with the numerical simulation, verifying the plausibility of the simulation. erefore, the distribution of deviatoric stress under the influence of multi-coal-seam mining is similar to the plastic zone in the surrounding rock of Tailgate 8709 in No. 11 coal seam. roof on the solid-coal side extends up to 4.0 m. ese corners are the key areas to control the surrounding rock of Tailgate 8709. In the support design process, it is necessary to strengthen the support such as anchor cables to the key control areas to prevent large deformation and damage of the surrounding rock in these areas [43]. e maximum distances of the peak zone of deviatoric stress are 4.0 m, 3.0 m, 4.0 m, and 2.5 m, respectively. e profiles of the peak zones of deviatoric stress and the plastic zone of Tailgate 8709 are shown in Figure 11. rough a comprehensive analysis of profiles of the peak zone of deviatoric stress and the plastic zone of Tailgate 8709, it is determined that the roadway roof and the shoulder corner of the roadway roof (in solid-coal side and coal pillar) have a large range of damage, which are beyond the support range of ordinary bolts. e roadway roof takes the combined support of Φ18 × 2000 mm bolts, Φ17.8 × 6300 mm single-anchor cables, and truss anchor cables. e singleanchor cable in the shoulder corner of the roadway roof on the solid-coal side is arranged at an angle of 20°on the inclined solid-coal side. e bolt parameters on both sides of the roadway are Φ18 × 2000 mm. Φ17.8 × 5000 mm and Φ17.8 × 4000 mm single-anchor cable are arranged antiasymmetrically on the coal pillar downward with an incline of 15°from the horizontal, and on the solid-coal side upward with an incline of 15°from the horizontal.

Mechanisms Using Combined Support by Bolts and
Anchor Cables. e roof and two ribs in Tailgate 8709 take Φ18 × 2000 mm thread steel bolts, which can connect the rock mass within the scope of shallow failure into a whole and enhance the integrity of the shallow roadway surrounding rock. e central anchor cable in the roof is arranged vertically, and the anchor cable in the shoulder corner of the roadway roof on the solid-coal side is arranged at an incline. At the same time, a single-anchor cable is inclined downward on the coal-pillar side and inclined upward on the solid-coal side passing through the peak zone of deviatoric stress. e high-strength and high pretension anchor cable can exert high-pressure stress on the roadway surrounding rock and using bolts can form a skeleton network structure in the pressure stress area. e anchor cable can tie the bearing structure formed by the bolts in the shallow surrounding rock to the deep stable rock strata. is makes the roof form a large-scale main bearing structure connected by the deep and shallow rocks and enhances the bearing capacity of the shallow surrounding rock of the roadway [44,45], thereby effectively inhibiting the separation of roof strata and maintaining the integrity of the bending deformation process of the roadway roof.

Supporting Mechanisms Using a Truss Anchor Cable.
A truss anchor cable is installed by two roof anchor cables at an incline at a place close to the two ribs. A certain preload is applied at the lower end of the anchor cable through a connector, so that the two anchor cables form a single bearing structure (Figure 11). e high pretension truss anchor cable system is used in the roadway roof to apply compound prestress to the regional roof fracture block and the rock mass above; this not only is beneficial to a roof in a state of multidirectional compressive stress but also improves the strength and deformation resistance of the coal and rock mass. It makes the surrounding rock within the anchorage range of the truss anchor cable consistent with the deformation of the upper rock mass and reduces the separation layer with the upper rock mass. At the same time, the truss anchor cable has a large length and strong shear resistance. e anchor cable in the two ribs of the roadway passes through the peak zone of deviatoric stress and the plastic zone and has a large range of action. It can effectively control the shear failure of the roof and the fracture of both sides of the layered roof structure and thus facilitates the balance of the overall structure of the anchoring zone [46].

Support Scheme of Tailgate 8709
e test roadway of this study is located at 350-470 m in Tailgate 8709. e roof is supported by threaded steel bolts, 18 mm in diameter and 2000 mm in length, with an interrow spacing of 1000 mm × 1000 mm. e roof bolts on the coal-pillar side are 750 mm away from the coal pillar and are installed at 70°from the horizontal. e bolts are connected by W-type steel strips of 3500 mm × 220 mm × 3 mm to form an integral bearing structure. e sizes of the steel pallets are 110 mm × 110 mm × 3 mm. High-strength and high pretension roof anchor cables, 17.8 mm in diameter, 6300 mm in length, and 1500 mm in array pitch, are installed at 55°f rom the horizontal on the solid-coal side to strengthen the support for the large-scale unstable rock in the roof corner. e roof support uses high-strength and high pretension anchor and truss anchor cables, 17.8 mm in diameter and 6300 mm in length. e single anchor cables are arranged between two rows of truss anchor cables with an array pitch of 4000 mm. e drilling depth of the single anchor cables, which are arranged vertically to the roof, is 6000 mm, while the drilling depth of the truss anchor cables, which have a pretension of at least 120 kN and are arranged at 20°to the vertical, is 5000 mm. In addition, the bottom span of the truss anchor cables is 2100 mm, and the distance between the borehole and solid coal rib is 1200 mm.

Rib Support.
e two ribs are supported by threaded steel bolts, 18 mm in diameter and 2000 mm in length, with an interrow spacing of 800 mm × 1500 mm. e distance between the top bolts of the two ribs and roof is 800 mm, and the top and bottom rib bolts are installed at 10°from the horizontal. High-strength and high pretension rib anchor cables, 17.8 mm in diameter and 5000 mm in length, are installed at 20°downward from the horizontal in the coal pillar, while the high-strength and high pretension rib anchor cables, 17.8 mm in diameter and 4000 mm in length, are installed at 15°upward from the horizontal on the solid-coal side. e rib-support anchor cables, with an array pitch of 1500 mm, could pass through the profiles of peak zone of deviatoric stress and the plastic zone, and the distance between the anchor cables (with steel pallets of 450 mm × 280 mm × 5 mm) and the roof is 1500 mm on the solid-coal side. e drilling diameter of all anchor cables and bolts in the above support is 28 mm, and the breaking load of anchor cables is 360 kN. e stereogram and planar graph of Tailgate 8709 support are shown in Figure 12.
To verify the accuracy of the plastic zone determined in numerical simulation and the practicability of the combined support technology for the surrounding rock around Tailgate 8709, a KDVJ-400 electronic drilling peep instrument [47][48][49] is used to observe the fracture zone development in the surrounding rock at the positions of the roof anchor cable in the roof corner, roof anchor cable on the coal-pillar side, and anchor cable in the coal pillar, as shown in Figure 12(b).
e carbonaceous mudstone at a depth of 0.2 m in the roof corner down the borehole is relatively fragmented, with massive longitudinal fractures, irregular fractures, and cavities. Furthermore, the opening degree of the fractures is relatively large, and the integrity of the rock is relatively poor, so that the hole can collapse easily in this region. At a depth of 0. To summarize, according to drilling peep instrument results, the plastic zone profile is located between the intact rock mass and the failure rock mass, and the length of the anchor cable is greater than the range of the loose failure rock mass in the shallow surrounding rock of the roadway. Hence, the accuracy of the plastic zone of the roadway surrounding rock in the numerical simulation has been validated by drilling detection, and the anchor cable can pass through the loose failure rock mass in the shallow surrounding rock of Tailgate 8709, effectively anchoring it to the deep stable rock strata.

Results and Analysis of Mine Pressure Observation.
In order to validate the control effects of combined support technology on the surrounding rock of Tailgate 8709 in No. 11 coal seam, the roof-to-floor convergences and rib-to-rib convergences of Tailgate 8709 have been monitored by an arrangement of monitoring stations; the observational results are shown in Figure 13. As shown in Figure 13(a), the roof-to-floor    convergences and rib-to-rib convergences of the Tailgate are 75 mm and 80 mm, respectively, without the mining-induced influence of the adjacent the N8707 coal face in No. 11 coal seam. When the adjacent coal face advance is 10 m ahead of the monitoring station, the roof-to-floor convergences and rib-torib convergences of the surrounding rock significantly increase. When the adjacent coal face advance is 10 m behind the monitoring station, the convergence rate of the surrounding rock is obviously reduced and the deformation slowly increases. e final roof-to-floor convergences and rib-to-rib convergences of Tailgate 8709 are 112 mm and 115 mm, respectively, and the deformation is controlled within a reasonable range. As shown in Figure 13(b), when the N8709 coal face advance is 30 m ahead of the monitoring station, the roof-to-floor convergences and rib-to-rib convergences of the surrounding rock gradually increase. When the coal face advance is 15 m ahead of the monitoring station, the convergence rate of the surrounding rock is obviously increased. is indicates that the Tailgate begins to be strongly influenced by the mining of the N8709 coal face, and the convergences of the surrounding rock linearly increase. e final roof-to-floor convergences and rib-to-rib convergences of the Tailgate are 172 mm and 198 mm, respectively. In addition, damage to support systems, such as the pullout and breakage of bolts and anchor cables, fails to take place in the field. Consequently, the combined support technology for Tailgate 8709 provides effective control of roadway surrounding-rock deformation under the disturbance of multicoal-seam mining.

Conclusions
In this paper, the deviatoric stress nephograms, stress curves, and plastic zone of the surrounding rock of Tailgate 8709 in No. 11 coal seam under the disturbance of the mining of the N8707 and N8709 coal faces in No. 7-4 coal seam and the N8707 and N8709 coal faces in No. 11 coal seam have been studied. e deflection angles of the peak zone of deviatoric stress after mining of the coal faces have been determined. e profiles of the peak zone of deviatoric stress and the plastic zone have been drawn. A combined support technology using high-strength and high pretension anchor cables and truss anchor cables has been proposed. e accuracy of the calculated plastic zone of the surrounding rock and the reliability of the combined support technology have been verified through drilling detection and mine pressure observation. e main conclusions are as follows: (1) After the mining of the N8707 coal face in No. 7-4 coal seam, the deviatoric stress of the surrounding rock of Tailgate 8709 in No. 11 coal seam has an approximately symmetrical distribution, which has low impact on the floor surrounding rock of the roadway. e mining of the N8709 coal face in No. 7-4 coal seam has a certain influence on the floor of Tailgate 8709, which results in the deflection of the peak zone of deviatoric stress in the shallow surrounding rock of the roadway. (2) e results of numerical simulation show that the peak zone of deviatoric stress of Tailgate 8709 in No. 11 coal seam continuously deflects under multi-coalseam mining. e peak zone of deviatoric stress in the roadway roof deflects by 34°, 13°, and 53°after the mining of the N8709 coal face in No. 7-4 coal seam and the N8707 coal face in No. 11 coal seam and in the advance 10 m of the N8709 coal face in No. 11 coal seam, respectively. is necessitates seeking a new effective bearing structure for the damaged surrounding rock. e supporting components such as anchor cables need to be anchored to the deep stable rock through the position of the peak zone of deviatoric stress. e maximum distance between the profiles of the peak zone of deviatoric stress and the plastic zone is not more than 1.0 m, and the plastic zone is large in the shoulder corner of the roadway roof in the solid-coal side and the coal pillar. (4) Based on the analysis of profiles of the peak zone of deviatoric stress and the plastic zone of Tailgate 8709 under multi-coal-seam mining, a combined support technology based on high-strength and high pretension anchor cables and truss anchor cables is put forward. Long, inclined, high-strength anchor cables in the shoulder corner of the roadway roof on the solid-coal side and in the coal pillar and short, inclined anchor cables on the solid-coal side are suggested as a means of strengthening the support. e results of drilling detection and mine pressure observation show that the deformation of the surrounding rock in Tailgate 8709 has been effectively controlled.
is research can provide a point of reference for roadway support under similar conditions.

Data Availability
Some data used to support the findings of this study are included within the article. Other data used to support the findings of this study are available from the corresponding author upon request.

Conflicts of Interest
e authors declare that there are no conflicts of interest related to the publication of this manuscript.

Authors' Contributions
Dongdong Chen, Shengrong Xie, and Fulian He conceived and designed the research. En Wang designed the model and the computational framework and wrote the paper. Long Wang was involved in the construction of numerical model. Qing Zhang and Xiaoyu Wu were involved in the numerical simulation. Zaisheng Jiang verified the analytical data. Yubo Li and Songhao Shi supplied the on-site data and verified the model in the on-site test.