Research on Stress Distribution Characteristics for Uniquely Shaped Roadway with Hard Roof in Steeply Inclined Coal Seam and Support Technology of Reducing Vibration Impact

This paper presents a comprehensive study of the stress distribution and stability analysis of a uniquely shaped roadway having a steeply dipping hard roof. The coal seam and its roof have a certain impact tendency, which is the internal condition of rock burst. The syncline tectonic stress causes the original rock stress to reach a higher level. The large amount of coal produced in the coal mine and the large movement range of the upper strata cause the huge mining additional stress around the stope. The impact load caused by “cantilever beam” fracture of hard roof can induce and strengthen rock burst. Its engineering geological setting encompasses the mining process and surrounding rock conditions of No. 6 Coal Seam in the 2130 coal mine of Xinjiang. Numerical simulations with theoretical analysis and ﬁeld measurements investigated a proposed new truss combined support scheme for implementation. A comparison was made of the diﬀerences in the state parameters of the road under the new and old support conditions. The application of the new combined support technology changed the form of the stress distribution around the road. Apart from the displacements of the two coal sidewalls, the new support system notably reduced the displacement of roof and ﬂoor by 67.8% and 83.6%, respectively. After the implementation of the new support scheme, the frequency of the original rock burst in the working face is greatly reduced, the surrounding rock control and ﬁeld application eﬀects also remained good, and personnel and equipment safety and production plan have a good guarantee.


Engineering Geology and Production Conditions
Steeply inclined coal seam refers to coal seams that have a dip angle greater than 45°. More than 50% of the coal mines in Southwest China operate in steeply inclined coal seams. Steep coal seams are widely distributed in many coal productions bases in Xinjiang, Gansu, Ningxia, Shanxi, Guizhou, Chongqing, Huainan, and so on. In China's total coal reserves and production over the years, steep coal seam accounts for less than 5%, but there are many production mines. According to incomplete statistics, steeply inclined coal seam mining is distributed in more than 100 production mines in more than 20 key mining areas in China, and 1/6 of the total number of mines in China is steeply inclined coal seam mining. With the focus of national resource development and coal mining shifting to the west, steep coal seam, which accounts for more than 50% of the number of mines in the west, has become an important mining object. Typical mines are Xinjiang Wudong coal mine, Gansu Yaojie No. 3 coal mine, Gansu Huating coal mine, etc. [1][2][3]. e main types of coal are coking coal, anthracite coal, and some scarce coal that are indispensable power fuel and industrial raw materials for national and local industrial development.
To adapt to the rapid growth of China's economy, the proportion of deep coal seams with a large dip angle in coal mining will increase year by year. in coal seams, large dip angles, and multilayering are the main characteristics of the steeply inclined coal seam. e stability of coal and rock masses on the mining road is the premise to ensure safe mining of the working face. Mine road openings usually have a rectangular cross section to meet the stringent transportation requirements set by the mechanized equipment operating in fully mechanized mining faces. Steep inclination angles lead to obtuse angles between the sides and top (roof ) of the road and between the roof and floor of the coal seam. Furthermore, the gravity force from the overlying coal and rock resolved along a steeply inclined bedding plane is more critically significant in considering the roadway failure than if the seams were horizontal or gently dipping. e critical areas for stability control of rock surrounding the roadway are also different, the uneven stress distribution caused by special-shaped roadway aggravates the frequency of rock burst, and, hence, it warrants a reasonable stability control method due to the prevailing asymmetric stress distribution in the surrounding rock.
With the continuous mining and consumption of mineral resources, the mining conditions of steeply dipping coal seams have changed from simple to complex. Mining of resources under such adverse conditions with a complex roof is worthy of scientific attention. e resource development processes in steeply dipping coal seam mining faces are more susceptible to complex mechanical problems than the horizontal (or gently inclined) coal seams. ese lead to frequent roof fall, rib spalling, and other related safety problems. Some scholars [4][5][6][7][8] believed that the deformation and failure of the rock surrounding roadways ( Figure 1) in steep coal seams have asymmetric characteristics. e stress distribution is asymmetric along the centerline of the road, which gives rise to coal sidewall failure and then leads to severe roof subsidence. Such asymmetric failures also include dislocation and subsidence of coal and rock mass at the shoulder corner. Its deformation, failure, and instability are a complex space-time problem. Some scholars [9][10][11][12] concluded that an unstable structure shows upward movement with tensional shear failure movement caused by the inclined gravity component for irregular quadrilateral road opening. Similarly, a downward trend of compressional shear occurs with the vertical component of the gravitational action resulting in a final cut or fall, respectively. e equations governing these two failure forms were established and presented together with the assessment criteria.
Taking the uniquely shaped roadway of 2130 coal mine in Xinjiang as the research background, this paper puts forward the common control method of roadway stability based on stress transfer from the roof and high wall vulnerable parts. It optimizes the construction scheme through theoretical analysis with numerical simulation and applies it to engineering practice successfully.

Engineering Background
e central coal seam of Xinjiang 2130 coal mine was the number 6 coal seam, which was the lowest minable coal seam in this mine. is coal seam dipped to the South, and the strike was near east-west. Observations from excavations at the low horizontal level drainage passage (+2047 m) and the 26,221-return air passage helped infer that the 3.7 m thick coal seam contained 0∼3 layers of gangue. It had a simple to medium structure, with a coal seam dip angle of 35°∼43°( mean dip of 39°). ere was a thin coal line in the number 6 coal seam roof, which led to poor contact between various strata. Roof falling and other accidents may have occurred during tunneling. Hence, the roof support work needed to be strengthened. e direct roof of the coal seam is comprised of coarse sandstone and glutenite. e bedding was the primary form of structure. e rock strength reduced on contact with groundwater when the rock effortlessly softened, causing block collapse. erefore, measures to strengthen the roof support were deemed extremely necessary. "Strike, long-wall mining method" was adopted on the 26,221-working face. e roof management mode was advanced preblasting to deal with the challenging and complex roofs. Notably, the inclined length of the working face was 148 m, and its dip angle was 38°∼44°. e cross section of the driving roadway was trapezoidal, with the taller wall being 4.5 m high and the height of the shorter wall being 1.5 m. e width of the roadway was 4.0 m. e roof structure of the roadway was simple, the direct roof was 2.0 m thick glutenite, and the main roof was 15.6 m thick medium sandstone. e activity at the mining face adopted the comprehensive mechanized backward mining method. It drove the return air passage and the transportation roadway along the roof, and the roof was not broken. e original trapezoidal roadway roof and the coal walls were supported by a conventional bolt (cable) monomer. e roadway roof was hard and had good integrity. Under the existing support conditions, the roof of coal seam has weak impact tendency, and roof fall and rib spalling frequently occurred during the tunneling process, which seriously restricted the safety production of mining enterprises. e most serious roof fall was caused by rock burst with a height of 26 m. e treatment of the roof fall resulted in 32 days shutdown of the mine, resulting in serious losses. Figure 2 illustrates the comprehensive geological stratigraphy of the coal seam, and Table 1 gives the physical and mechanical parameters of coal and rock strata.

Stress Analysis of Roadway Roof.
According to the actual geological conditions of the 2130 coal mine, establishing a reasonable mechanical model was the basis of the analysis. e roadway roof was simplified as a simply supported beam structure [13]. Based on the small deflection and elastic thin plate theory, the working face model is established, and the mechanical analysis of the upper main roof of the stope is carried out [14]. Unsymmetrical deformation mainly occurs near the weak structural plane and the intersection of roadway section and surrounding rock inclination direction. Based on this, the "unsymmetrical high prestress pressure relief coupling control technology" is proposed, and its analysis and industrial test are carried out. e practice shows that the asymmetric coupling support technology can not only effectively solve the problem of differential deformation of surrounding rock, but also ensure the coordinated deformation of support structure and Medium-grained sandstone, subangular to angular, and hard It contains quartz crystal, hard, and it is hard and easily falls off.
The coal seam structure is simple to medium, uneven, and banded in structure, and endogenetic fractures develop.

Dark gray carbonaceous mudstone and dense
Medium thick bedded argillaceous siltstone is easy to soften with water and filled with argillaceous cement.    Shock and Vibration surrounding rock, so as to improve the overall stability of the roadway, greatly reduce the repair rate of the roadway, and save the resource support cost [13][14][15][16][17][18]. e shorter sidewall was simplified as a fixed hinge support A, and the taller sidewall was simplified as a movable hinge support B. Figure 3 shows the established mechanical model. In Figure 3; F 1 and F 2 are vertical bearing reaction forces, while F 3 is the horizontal force at the fixed hinge bearing end. While q is the stress of the upper load acting on the roof, ρ is the rock mass density. X is the horizontal distance between the points A and B and l is the width of roadway. e acceleration due to gravity is g and α is the dip angle of the coal seam.
In an irregular-shaped roadway, with an increased coal seam dip angle, the height of the higher sidewall is increased. e combined action of vertical and inclined stresses caused the two coal sidewalls to be more prone to different levels of damage. e damage degree and scope of the high wall were much larger than those of the low wall, further aggravating the roof settlement and increased risk of roof accidents. Under the combined action of mining depth increase, driving influence, and other factors, the load on the upper part of the roof showed a dynamically increasing trend, and the vertical deformation of the roof continued to increase: According to the vertical-static equilibrium condition, the roof deflection curve equation was obtained as follows: where E is the elastic modulus of roof rock mass, I is the bending moment of inertia of roof, and ɷ represents the deflection of the roof. Using the lithology parameters and formulas (1) and (2), the deflection relationship between different upper load and roof subsidence was obtained. e roof subsidence was positively correlated with the upper load, and the consequent deflection curve was in the shape of a net pocket. Due to the stress asymmetry on the roof from the uniquely shaped roadway, the bottom of the deflection net pocket curve deviated from the center of the roadway and was closer to the low sidewall. e bottom of the net pocket was the maximum tensile stress on the roof, which was, therefore, the crucial monitoring and protection area. Subsequent numerical simulation and field monitoring also verified this point.

Stress Analysis of Roadway Coal Wall.
Under the influence of tunneling and roof stress transmission, the coal walls of the roadway were easy to form endogenous cracks, and improper support of the roadway facilitated roof fall accidents. Under the influence of vertical stress, the coal walls were prone to slip failure. According to the analysis of the roadway side slip mechanical model [13] and field practice, the middle and upper parts of the higher coal wall were the most vulnerable areas. If the two sides of the coal wall structure were intact, it could support the roof load. If the two coal sidewalls were impaired, the roof could sink as a whole, and the roadway structure will be significantly damaged, which led to roof fall and rib spalling accidents. e following three technical measures enhanced structural stability and changed the stress state of the roadway. (1) e support anchor point was located in the stable rock layer to bear the roof pressure and to reduce the pressure transfer load of the roof onto the two coal sidewalls.
(2) We changed the original support form of roadway, increasing the truss connection between anchor cables to improve the shared bearing capacity between the support structure and the surrounding rock. (3) As per the roadway condition, the pressure relief groove was designed to minimize stress concentration. Figure 4 shows the overall map of the field roadway under the new truss anchor cable support structure. e roof and two sides of the coal wall are complete without obvious damage. After the implementation of truss support structure, the structural integrity of roadway is good, the deformation of roof, floor, and two sides of roadway is small, and the rock burst phenomenon is significantly reduced, which meets the needs of safety production. e basic idea of roadway safety control was to transfer the stress of roof and coal wall to the deep coal and rock mass, reduce the stress damage on the coal sidewalls, and improve the safety of the roadway construction, improving the excavation n efficiency and mine production efficiency.

Model Establishment and Parameter Selection.
Based on the geological conditions of the roadway, the numerical model of coal and rock strata was established with the FLAC3D software, and the Mohr-Coulomb failure criterion was adopted. Considering the coal seam dip angle and model block size, the final modeling size was 30 m × 35 m × 0.1 m (long × width × high). e block size of the model element e whole 3D model was divided into 126,996 units and 218,675 nodes. e roadway section was an irregular trapezoid, and the roadway size was (high side 4.5 m + low side 1.5 m) × 4.0 m (height × width). Figure 5 illustrates the establishment of the numerical model. Combined with the actual situation of the driving roadway, the vertical displacement of the bottom boundary of the model was set to zero, and the horizontal displacement of the front, back, left, and set right boundaries of the model was set to zero. Considering the influence of self-weight stress of strata, (2.7 × 5.3 � ) 14.31 MPa stress was applied to the overlying strata as the self-weight compensation load uniformly applied to the upper interface of the model.
Numerical simulation of roadway without support, original roadway support, and new truss structure support three forms. In the study of microseismic events and numerical simulation of rock burst mechanism in underground coal mine roadway, some inspirations have been obtained from the article [19,20].

Analysis of Surrounding Rock Stress State.
ere was no other working face around the driving roadway during the period of this research. Hence, the analysis did not consider the influence of mining factors. To carry out a vivid comparative analysis of the stress state in the surrounding rock under different support forms, this research considered three different roadway states, including bare roadway without support, original support (without truss structure), and truss anchor cable support system. Figure 6 presents the numerical simulation results showing the stress profiles at the observation points on the roadway roof when under the three forms of support systems mentioned above, and the vertical stress distribution along the coal seam strike roadway section is shown in the figure. e stress distribution of the roof was approximately arched. e maximum tensile stress on the roof was 2.5 MPa, which did not reach the ultimate tensile strength of 3.6 MPa. e roof structure of the roadway was still intact and transferred the overlying load to the two coal sidewalls. e simulation results displayed in Figure 6 indicate that the range and intensity of the rock stress area at the top of the roadway were progressively increasing from case (a) with no roadway support to case (b) that is the original support state to case (c) that is the new truss anchor structure support state. Simultaneously, the rock stress at the top of the roadway was gradually decreasing, with the color becoming lighter, thus, indicating that the new support structure had successfully transferred the roadway confining pressure to the deeper rock area and reduced the roof pressure and the damage degree on the shorter coal sidewall.

Stress Analysis of Roof Strata.
e objective in the new support design was achieved.
Vertical stress nephograms are shown in Figure 6 for the three kinds of roadway support forms. It is noteworthy that the maximum tensile stress of the roadway roof without support was 32.4 kPa, and the maximum compressive stress was 35.8 MPa. e maximum tensile stress of the roadway roof was 57.0 kPa, and the maximum compressive stress was 30.7 MPa. e maximum tensile stress of the roadway roof under truss support structure was 41.2 kPa, and the maximum compressive stress was 29.5 MPa. In contrast, on the premise of ensuring a good overall structure in the surrounding rock, the tensile stress of truss support structure was 27.7% lower than that of the non-truss anchor cable support structure. e maximum tensile strength of roof rock in the natural state was 2.1 MPa. Although the roof without support or truss support did not exceed the ultimate limit, there was no apparent damage. e three different support systems gave differing stress distributions (Figures 6(a)-6(c)) in the surrounding rock. e red range of tensile stress around the roadway decreased in turn indicating that the tensile stress state of the roadway roof and the surrounding rock was weakening. is trend was extremely beneficial to the protection of the roof and surrounding rock. e roof stress in Figure 6(c) was within the ultimate tensile strength, and the distribution of deep compressive stress was the smallest among the three figures.
is indicated that the stress in the shallow area around the roadway was controllable. e transfer stress in the deep area was reasonably distributed, and the support structure made Truss anchor structure at roof Truss structure at high side Truss structure at low slope Shock and Vibration the overall stress of the roadway and surrounding rock more uniform, which was the best state desired for the actual construction. Figure 7 displays the maximum subsidence displacement diagrams with the three different roof supports. According to these curves of the roadway roof subsidence, the maximum vertical displacement of the roof was 50 mm under the condition of no support, 16 mm under the condition of original support, and 15 mm under the state of new truss support. It showed that the roof structure was completely intact, the stress of overlying strata and original rock was not reflected on the roof, and the stress on the roof was effectively transferred to the coal sidewalls. ese results show that the maximum compressive stress is 1 MPa without support, 2.0 MPa without truss cable support, and 2.4 MPa with truss structure support. is effective stress variation trend indicated that the roof stress state change from a tensile stress state to a compressive stress state was becoming more substantial. e compressive stress was far less than the limiting compressive strength value of the rock. e surrounding rock structure remained good and unchanged.
On the close comparative study of the effective stress history curves for the roof, it was evident that the effective    Shock and Vibration stress in Z direction for the roof with no support was 1.0 MPa. When the roof had the original support and under the new truss anchor supports, the effective stress on the roof was 1.5 MPa and 1.8 MPa, respectively. e Mohr-Coulomb failure criterion points out that the smaller the effective stress of rock is, the greater the fluid pressure is, and the easier it is for the rock to develop tensile failure. is confirmed that, of the three methods compared, the new truss support method provided an enhanced positive role in the roof control, with the overall state of the roof remaining more stable, less prone to be impaired. us, the new truss and anchor support was far more effective in maintaining the roof's integrity and roadway safety. Figure 9 compares the roof displacement profiles resulting from the three different support systems studied. It presents that the roof displacement was reduced from 4.5 cm in no support to 1.9 cm in the state of no truss anchor support. e introduction of the anchor support to the truss structure reduced the roof displacement further to 1.8 cm. Notably, the roof displacement of the truss structure did not give an extremely noticeable improvement compared with that of the nontruss anchor cable support structure system. is was because the truss support structure had a better embodiment within the deeper stress transfer and distribution, giving the overall protection to the coal sidewall structure.

Analysis of the Stress Levels on the Two Coal Sidewalls
(1) Following the above mechanical analysis of the structure, Figure 10 illustrates the changes in the distributions of the plastic zone in the surrounding rock. It became evident that, under the action of vertical stress, the joint gap of the high coal sidewall produced staggered shear, and the tensile action made it reach towards the compression shear failure limit. e stress in the middle and upper areas of the high coal sidewall was concentrated, and this serious damage extended to the deeper part of the coal body. e maximum failure depth in the high coal sidewall was 2.65 m. erefore, the failure probability in the high sidewall was much greater than that in the low coal sidewall. Step ×10    Step Step ×10  Step Step ×10  Step  (2) e smaller height of the short coal sidewall and its relatively stable mechanical structure made the failure of the coal body not so apparent with the action of stress. e maximum failure depth in the short sidewall was 0.6 m, which amounted to only 22.6% of the maximum failure depth of the high wall.

Analysis of the Changes in the Plastic Zones
(1) Considering the condition of the overall roadway structure, under different support conditions, the high coal sidewall was severely impaired by stress; the plastic zone was large, and the stress damage went deep into the coal wall. e anchor point within the support structure scheme needs to reach the stable rock layer before the support action can take effect. Increasing the intensity of support is the only means to control the roadway deformation. To enact this, the cost is high, and the controlling effect from the rock surrounding the roadway is poor. (2) e results in Figure 10 show that the height of short coal seawall with a low slope was small, its structural stability was strong, the damage of the coal sidewall under stress was light, and consequently, the damage range of the plastic zone was generally tiny. (3) e plastic zones for the roadway without support were concentrated around the roof and the coal sidewalls. e damage around the roadway was severe. It slightly reduced the stress regime in the rock surrounding the roadway (without truss anchor support), but the damage to the upper corner of the roof and the high sidewall was still apparent.   However, the new truss support scheme significantly reduced the plastic zones on roof, and the plastic zone in the deep parts of the surrounding rock, especially in the deep part of the floor rock, was greatly increased. e support structure achieved the intended objectives of the original design of transferring the stress around the roadway to the deeper part of coal and rock to protect the safety of the driving roadway effectively.

oughts on Control of the Rock Surrounding the Roadway.
Mechanical structure analysis and the corresponding verification with numerical simulation facilitated identifying the areas that were prone to failure. ese were the middle and upper portions of the roadway roof and the high wall. e combined support scheme strengthened the support for these critical parts of the roadway. e combined support technology from "truss anchor cable + W steel belt + anchor rod + anchor mesh" was a further asset to the scheme. e pressure relief groove was excavated on the two sides of coal wall in time to avoid high stress concentrations. e roadway truss combined support technology formed an effective linkage structure to minimize the stress and deformation on the roof and the two sidewalls [21][22][23][24][25].
e supporting principle was mainly to form an encircling force structure in the zone of high-stress risk through the anchor cable and the connector anchored into the stable position of the roof. As a result, the bending roof subsidence and the broken gangue were mutually compressed and self-stabilized, forming a joint support structure with the coal and rock mass. is enabled the transfer of the stress, which initially acted on the two coal sidewalls to be directed to the stable primary roof rock. It then significantly reduced the pressure and deformation degree between the direct roof and coal sidewalls. Figure 11 illustrates this further through the schematic diagram for the combined support from the joint truss structure, and details of the truss anchor cable connector are as shown in Figure 12.

Combined Support Scheme and Key Technical Parameters
e original roadway support was based on a negative and passive support technology. e combined support scheme mainly improved the support strength by increasing the support density. In the process of working face excavation, the abnormal mine pressure behavior characteristics of the roadway are as follows: bolt cable fracture and failure, the roadway roof subsidence, two sides approaching towards each other too much, the frequent occurrence of roof and coal sidewalls caving accidents, and the roof caving height varied from 1 to 4 m. It confirmed that the effect of simply increasing the support density was not the ideal solution.
e new support adopted the "truss anchor cable + W steel belt + anchor rod + anchor net" in the roof and high sidewall. is was a kind of positive and active technical support. Figure 13 shows the layout of the combined support. e specific supporting characteristics were as follows:

Bolting Parameters.
Compared with the original support system, the new truss support method mainly increased the truss anchor cable structure reducing the anchor cable support density support and did not change the spacing arrangement comparison is shown in Figure 13. e high sidewall bolt was 18 mm in diameter. is ribbed steel bolt length was 2300 mm, with an anchorage length of 1200 mm. e pretightening torque was to be not less than 250 Nm. e distance between bolts and rows was 1000 mm * 1000 mm for the original support system. e new support system changed the corresponding spacing to 1200 mm * 1000 mm. e comparison of new and old support parameters is shown in Table 2.

Parameters of the Anchor Cable Support
(1) e roof truss anchor cable was 17.8 mm in diameter and 6900 mm long. It was a high-strength anchor cable with a resin anchored cable length of 4900 mm deep into the roof section. An 1800 mm anchor section length, with a further 200 mm length, was reserved at the end of the anchor cable to overlap with the anchor cable connector. e anchor cable pretension was not to be less than 140 kN. e roof truss anchor cable row spacing was 4000 mm, with the high sidewall anchor cable and the roof subtending an angle of 34°. e angle between the anchor cable and the roof at the low sidewall was 83°. e original support system had only one single roof anchor cable and had no truss anchor cable.
(2) e anchor cable on the high sidewall truss was 17.8 mm in diameter and 6500 mm long. It was a high-strength anchor cable with a resin anchored cable length of 4600 mm deep into the roof section. An 1800 mm anchor length, with a further 200 mm length, was reserved at the end of the anchor cable to overlap with the anchor cable connector. e anchor cable pretension was not to be less than 140 kN. e row spacing of the anchor cables on the high sidewall truss was 4000 mm. e angle between the anchor cable to the roof and the coal wall was 74°, while the angle between the partial floor and the coal sidewall was 84°. ere was no high sidewall truss anchor cable used in the original support system. (3) We installed a single anchor cable between the two rows of truss anchor cables on the roof. is single    Shock and Vibration 13 anchor cable was 21.8 mm in diameter, and the anchorage length of the high-strength anchor cable was 1800 mm. e original roof support system also had a single anchor cable.

Analysis of the Effects from the Combined Support Systems.
Dynamic measuring points were arranged along the roof and coal sidewalls at 20 m ahead of the working face. Four measuring points were in each group and arranged three groups making a total of 12 measuring points. e observed data from such measuring points were counted and presented in Figure 14. is figure compares the coal and rock mass deformation under the new and old support systems. Figure 14 shows that the maximum roof pressure under the original support condition was 44 MPa, and that under the new support condition was 27 MPa, which amounts to 61.4% with that from the original support system. e maximum displacement of the roof and floor under the original support was 25.5 cm. In contrast, the maximum displacement of roof and floor under the new support was only 8.2 cm, which amounts to 32.2% of that from the original support system. Under the condition of the original support, the maximum approach of the two sides was 55 cm. In contrast, under the needs of the new truss combined support system, the maximum approach of the two sides was only 19 cm, which amounts to only 16.4% of that from the original support system. e overall effects on the roadway control were remarkable with the adoption of the new support system. According to the online monitoring system of mine pressure and the statistics of field production data, after the implementation of the new truss anchor cable, the stress of surrounding rock is stable and far below the bearing limit of roof, and the number of rock bursts in the mine is significantly reduced, from the original average of twice a month to zero. e stress stability of surrounding rock represents that the probability of impact is extremely small, which is verified by the field production. According to the comparison of the frequency of roof falling and spalling and the size of spalling in Table 3, it can be seen that the monthly average spalling frequency of the working face is reduced from 23 times to 4 times, the spalling frequency is reduced by 82.6%, and the size of spalling and spalling is also greatly reduced. is also proves that the implementation of the new truss weakens the stress of surrounding rock and reduces the impact tendency of coal mine. During the whole mining period of the working face, there is no serious injury and impact accident of personnel and equipment, which effectively ensures the safety in production.

Conclusions
e following conclusions can be drawn from the research carried out on the comparative performance study on the roadway control support systems: (1) e new type of combined truss support on the roof and high coal sidewalls presents circular compression support forming combined structural support with the body, and coal rock mass bearing together. e above facilitated a noticeable control effect on the stability of the roadway.
(2) With the implementation of the new support technology, the maximum distance between the roof and floor was 8.2 mm, and the maximum distance between the two coal sidewalls was 19 mm, which amounts to 67.8% and 83.6%, respectively, lower than that from the original support system, and the spalling frequency is reduced by 82.6% after the implementation of the new truss support measures. e above fully meets the needs of safe coal production.
(3) e anchorage position of the support body must break through the shallow coal and rock mass and go deep into the stable rock stratum to play the desired supporting role. Increasing the support density alone will not effectively improve the roadway control effect. (4) e application of new combined support technology with the pressure relief groove was designed to enhance the control effect of the roof and the two coal sidewalls and excavated according to a situation imposed on a roadway to prevent stress accumulation. However, there is still room for optimizing technical parameters in the support scheme to carry out more in-depth research in the future.

Data Availability
e data used to support the findings of this study are available from the corresponding author upon request.

Conflicts of Interest
e authors declare that they have no conflicts of interest.