A novel non‐pillar coal mining technology in longwall top coal caving: A case study

In longwall top coal caving (LTCC) mining methods, problems such as large deformations, loss of coal resources, and excavation relay out of balance are often encountered with the increased production in single mining panels. To address these problems, a novel non‐pillar coal mining technology in LTCC was used based on the equilibrium mining theory. This approach could cut off the stress transmission, eliminate the coal pillar, and retain the gob‐side entry. In this technology, the entry is first excavated along the roof of the coal, and then four techniques are applied to control the surrounding rock of the retained entry. The constant resistance and large deformation anchor cable are used to maintain the roof stability. The directional roof split blasting technique is adopted to separate the roof between the entry and gob. The blocking‐gangue support system (BGSS) is used to form the gangue rib lagging the mining panel. The temporary support system is applied to prevent the roadway from being affected by the dynamic pressure and prevented the BGSS from falling. Finally, the field monitoring results indicate that the surrounding rock of retained entry is stabilized at 200 m behind the working face. The deformation of retained entry is within a reasonable range at 180–190 m behind the mining panel. The results provide an important reference for the thick coal seam non‐pillar coal mining technology, specifically using this technology in LTCC panels.


| INTRODUCTION
Energy is the basis for the progress of human civilization and is related to human survival and development. Global energy demand continues to grow, at least for a period, driven by increasing prosperity and living standards in the emerging world. 1 Coal is the basic energy source for ensuring energy supply and plays a significant role in maintaining the international energy structure. The LTCC mining method is one of the most important coal mining methods; its progress has greatly promoted the production of coal after the end of the seventeenth century. 2,3 In the past decades, the LTCC mining method has been widely employed in Chinese coal mines, especially for extracting coal seams with a thickness of more than 4.5 m. 4,5 However, with the output increases per mining panel, there are some problems when using the LTCC mining method: (1) large deformation of mining entry; (2) methane control 6 ; (3) low recovery ratio 7 ; (4) excavating relay out of balance 8 ; and (5) spontaneous combustion of residual coal in gob. 9 Many scholars are dedicated to research on controlling the roadway large deformation around the world. Kang et al. proposed a reinforcement supply method including high pretension intensive bolts and cables after high-pressure grouting to control the roadway's large deformation. 10 Wen et al. first provided a method for determining the reasonable support strength of the bolt to control the roadway large deformation based on the theory of stress gradient of the surrounding rock. 11 Yang et al. proposed coupling support of a double-layer truss to control the deformation failure of the extremely soft rock roadway. 12 However, these support methods only support the surrounding rock of the roadway and cannot increase the resource recovery rate and improve the roadway excavate ratio.
Therefore, Zhang et al. proposed to use the gob-side entry retaining (GER) technology in the LTCC panel to cancel the coal pillar, improve the resource recovery rate, and decrease the excavating ratio. 13 However, GER technology has not changed the roadway stress environment, which is the problem of stress concentration caused by the coal pillars. In addition, as the production of LTCC panels increases, it requires a larger cross section of roadway, and the mining height and daily footage of panels also increase. 6 It will lead to the stress concentration on the packing wall, and the packing wall and gob-side entry deform and collapse easily. Therefore, the gob-side packing wall must have a larger width, height, and higher strength. But the speed of construction of the gob-side packing wall seriously influences the mining efficiency. 14,15 Furthermore, although the coal pillars were eliminated, another packing wall is still retained. Therefore GER technology could not eliminate the geologic disasters caused by the stress concentration of coal pillars, such as rockburst, gas outbursts, and serious deformation of surrounding rock. [16][17][18] To completely eliminate the coal pillars, He first proposed the "equilibrium mining theory (EMT)" and formed a novel non-pillar coal mining technology on the basis of EMT. 19 This technology proposes to automatically form a gob-side entry by utilizing the ground pressure and eliminating the coal pillars or packing walls between two adjacent panels. 20 In this technology, the stress transition between the roof of the roadway and the goaf is first cut off by the directional roof splitting blasting (DRSB) technique ahead of the mining panel, and then the gob-side entry is automatically retained by using the ground pressure. 21,22 Moreover, due to the bulk factor property of gob materials, the collapse gangue behind the panel could completely fill the gob and support the overlying strata. The gob materials can form gangue ribs of the retained entry. 23,24 This technology could reduce the 50% roadway drivage ratio and increase the recovery rate of coal resources to 100% in the stope. 25 The first industrial experiments of this novel nonpillar coal mining technology in the Baijiao Coal Mine were successful in 2009. 26 In the next few years, through experiments in the Jiayang Coal Mine, and Hecaogou No. 2 Coal Mine, the possibility of this technology in thin coal seams was verified. [27][28][29] The adaptability of this technology in middle thick coal seams was demonstrated by the test of the Chengjiao Coal Mine, Ruineng Coal Mine, and Halagou Coal Mine. [30][31][32] The feasibility of this technology in thick coal seams was examined by the experiment in the Ningtiaota Coal Mine, Baoshan Coal Mine, and Hongqinghe Coal Mine. 23,24,30 In recent years, this non-pillar coal mining technology has been promoted and applied in many coal mines and panels in China as shown in Figure 1. However, the existing research is concentrated on one-pass cutting coal mining, and the mining height of the test mining panel is less than 4.5 m.
When the mining height is greater than 4.5 m, the larger mining height will cause more serious deformation of the roadway surrounding rock and more serious loss of coal resources. In addition, whether the mining roadway is excavated along the roof or floor of the coal seam, the top coal or the bottom coal will cause the roadway surrounding rock more difficult to control. Moreover, there are great differences in the mining process and ground pressure law between the LTCC panel and the one-pass cutting coal mining panel. Therefore, to improve the resource recovery rate of the LTCC panel and reduce the large deformation of roadway surrounding rock, the authors propose a novel non-pillar coal mining technology in LTCC panels.
This paper studies the non-pillar coal mining technology in LTCC panels based on the Shenlei coal mine. Considering the characteristics of mining roadways the top coal is loose and easy to collapse. Meanwhile, the gangue rib of retained entry is difficult to control under the impact of the collapsed gangue in the gob. First, the entry is excavated along the roof of the coal seam to facilitate support retained entry. Second, using four key techniques to control the surrounding rock of retained entry in LTCC panels including constant resistance and large deformation (CRLD) anchor cable, DRSB technique, blocking-gangue support system (BGSS), and temporary support system (TSS). Finally, a novel non-pillar coal mining technology in LTCC panel is proposed based on the above techniques and applied in field engineering.

| Mining specifications
The Shenlei Coal Mine is located in Heshun County, Jinzhong City, Shanxi Province, China, as shown in Figure 2. Its production is designed to be 0.9 Mt/a. The No. 15 (16 m). As shown in Figure 2, the generalized stratigraphic column of the 150202 mining panel, is drawn based on the drilling cores in the roof and floor of the tailgate.
The strike and dip length of the 150202 LTCC panel is 1240 and 150 m, respectively, as shown in Figure 2. The surrounding rock of the 150202 tailgates is severely deformed during the excavation process, which seriously affects the speed of roadway excavation because the 150201 working face adopts the traditional LTCC mining method and leaves a coal pillar with a width of 20 m, as shown in Figure 3.
The fundamental reason for the large deformation of the 150202 tailgate is the stress concentration of surrounding rock caused by the coal pillars. To eliminate coal pillars and stress concentration, the 150202 LTCC mining panel was selected for experiments to demonstrate the possibility of the novel non-pillar coal mining technology in LTCC panels. Considering the production schedule of the Shenlei Coal Mine, first, the 150202 LTCC mining panel is mined 240 m by the traditional LTCC mining method, and then the experiment parts of non-pillar coal mining technology are 1000 m, as shown in Figure 2. To facilitate support of the retained entry, the 150202 headgate is designed to excavate along the roof of the No. 15 coal seam and the first five hydraulic supports  of retained entry side gradually increase the caving height.

| Equilibrium mining theory
In 1960, Qian first proposed the "masonry beam theory" and conducted a comprehensive study of the stress transfer structure of the overlying strata above the gob. 33 In 1980, Song proposed the "transferred rock beam theory", which further explained the transmission path and stress distribution of the overlying strata. 34 Wen first constructed a spatial structure model and proposed the conditions and control measures for coal mine disaster accidents. 35 However, as the mining depth increases, the mine pressure becomes more serious, and the large deformation and damage problem becomes more difficult to control. Therefore, to reduce the mine pressure disaster from the source, he put forward the EMT and proposed to fill the gob with collapse gangues and automatically form the roadway. 36 A large area of gob ∆V m will be left after the working face advanced, as shown in Figure 4A. As the gob area increase, the overlying strata above the gob collapse, and the fractures gradually extend to the ground surface, eventually, causing surface subsidence. When the caving rock is in the process of compacting, the overlying strata above the gob gradually formed three zones (caving zone, fractured zone, and bending zone).
To describe the damage changes in different zones of the overlying strata above the gob area, we define three damage variables, K 1 , K 2 , and K 3 . K 1 is the surface subsidence damage variable, which is calculated by dividing the volume change of the surface subsidence ∆V S by the volume change of mining ∆V m (Equation 1). K 2 is the fracture damage variable, which is calculated by dividing the volume change of the fractured band ∆V f by the volume change of mining ∆V m (Equation 2). K 3 is the caving damage variable, which is calculated by dividing the volume change of the caving band ∆V C and the volume change of mining ∆V m (Equation 3), as shown in Figure 4B. Although these three damage variables are always changing with different mining methods or geological conditions, their sum remains unchanged at 1 (Equation 4). In the traditional LTCC mining method, the value of K 1 can be obtained through field measurement, but K 2 and K 3 cannot be measured. Therefore, the ground subsidence cannot be reduced by increasing the fracture damage variable and collapse damage variable; this mining method is an unbalanced mining system.
However, if the height of the caving zone can be obtained and further control the height of the caving zone by blasting, then Equation (4) can be solved. In non-pillar coal mining technology, the expansion property of gob materials is utilized to increase the height of the caving zone by using the DRSB technique. When the ∆V C is equal to the ∆V m , surface subsidence will be eliminated, ∆V = 0 f and ∆V = 0 S , as shown in Figure 4C. Therefore, this mining technology is a balanced mining system. The most reasonable roof cutting height H s can be calculated by Equation (5) and the average bulking factor of the overlying strata K p can be measured using Equation (6).
where K p is the average bulking factor of the overlying strata, H s is the DRSB height (m), S is the mining area (m 2 ), K 0 is the initial bulking factor of the overlying strata, α is the fitting coefficient, and t is the time.

| A reasonable roadway layout in LTCC
Compared with the thin coal seam mining panel, the LTCC panel in the thick coal seam needs to consider the roadway layout. For the GER technology in thick coal seams, the mining entry is designed to excavate along the floor of the coal seam because it is necessary to maintain the stability of the packing wall and prevent it from dumping, as shown in Figure 5A. 6,37 However, when excavating along the floor, the top coal above the roadway will experience multiple mining disturbances; the top coal is difficult to control, and roof caving occurs easily. In addition, due to the large mining disturbance in the LTCC panel, the BGSS behind the panel easily dumps when it is disturbed by the collapse of the gangue. However, if the roadway is excavated along the roof of the coal seam and the triangular bottom coal is left on the floor of the retained entry side, the impact disturbance of the collapsed gangue will be significantly reduced. Therefore, excavating along the roof of the coal seam is a reasonable roadway layout when using nonpillar coal mining technology in LTCC panels, as shown in Figure 5B.

| The principles of non-pillar coal mining technology
The non-pillar coal mining technology adopts the DRSB technique, which can not only increase the height of the caving zone for reducing the surface subsidence, as shown in Figure 4C but also cut off the stress transmission between the roof of the gob area and the retained entry. Then the collapsed gangue will fill the gob area and eventually support overlying strata. Therefore, the retained entry is in the pressure relief zone. When the non-pillar coal mining technology is applied in LTCC panels, the design of the LTCC panel and entry formation process are as follows: first, the entry is excavated along the roof of the coal seam, as shown in Figure 6A. Second, using CRLD anchor cable to reinforce the support strength of the entry roof that needs to be retained. Moreover, it is necessary to use the anchor cable to reinforce the support strength of the coal-mass rib to prevent the rib from falling, as shown in Figure 6B. Third, DRSB technology is performed to split the gobside roof of the retaining entry ahead of the mining panel, as shown in Figure 6C. Fourth, the overburden above the gob influenced by DRSB will cave and form a gangue rib of retained entry under the support of BGSS and the TSS technique, as shown in Figure 6D. Finally, as the caving gangue is gradually compacted and stabilized, the TSS is withdrawn, and the retained entry will be stable.
F I G U R E 5 View of a reasonable roadway layout in longwall top coal caving: (A) excavate along the floor of the coal seam when using gob-side entry retaining technology, and (B) excavate along the roof of the coal seam when using non-pillar coal mining technology.

| CRLD anchor cable support
The retained entry will experience many disturbances, resulting in large deformations after using non-pillar coal mining technology in LTCC panels. However, the elongation of anchor cable made of traditional steel strands is less than 5%, which cannot meet the deformation requirements of the roadway surrounding rock and is unable to adapt dynamic pressure shocks in the process of retained entry. Therefore, He et al. invented the CRLD anchor cable. [38][39][40] It can accommodate the large deformation of the roadway surrounding rock and adapt to dynamic pressure shocks caused by collapse gangue.
The CRLD anchor cable is mainly composed of four parts: the constant resistance body, steel strand, face pallet, and anchor cable cone lock, as shown in Figure 7. The constant resistance body is the key structure of the CRLD anchor cable, which is mainly composed of the sleeve pipe and cone lock. The constant resistance of the CRLD anchor cable comes from the friction of the cone lock in the sleeve pipe because the sleeve pipe provides a track for the movement of the cone lock. By installing the constant resistance body on the rear part of the ordinary steel strand, the cone lock is used for fixing the free end on the rear side of the constant resistance body to provide stress release and constant resistance. Since the anchoring end of the CRLD anchor cable is fixed in the main roof, only the cone lock at the free end can slide. Therefore, when the roof of the retained entry is deformed under mining disturbance, the cone lock at the free end slides friction in the sleeve pipe, resulting in constant resistance, and the deformation value of the anchor cable is determined by the length of the sleeve pipe. The CRLD anchor cable realizes the negative Poisson's ratio effect through the structural yield of the constant resistance body. At present, the CRLD anchor cable can reach a maximum tensile length of 1000 mm under a constant resistance load of 130-850 kN. 41 Figure 7 shows the support principle of the CRLD anchor cable. Figure 7A shows the original state installation of the CRLD anchor cable to support rock mass. When installing the CRLD anchor cable, first use a resin capsule to fix the steel strand in the main roof, then install a constant resistance body on the rear free end of the steel strand and use a cone lock and face pallet to fix the constant resistance body. In this stage, the elastic deformation is small, and it mainly occurs on the steel strand, cone lock, and constant-resistance body. When the axial force of the CRLD anchor cable reaches the constant resistance P 0 , the cone lock will slide into the sleeve pipe of the constantresistance body, as shown in Figure 7B. The constant resistance structure will undergo plastic deformation. The constant resistance P 0 of the CRLD anchor cable is influenced by the elastic characteristics of the sleeve pipe, the structure of the cone lock, and the friction properties of the contact surface between the cone lock and the sleeve pipe. Therefore, when the external load of the anchor cable exceeds its constant resistance P 0 , the lock will produce a relative axial displacement in the sleeve pipe and absorb a lot of deformation energy generated by the overlying strata. The maximum axial deformation of the CRLD anchor depends on the length of the sleeve pipe. Figure 7C shows the maximum axil deformation stage of rock mass supported by the CRLD anchor cable.

| DRSB technique
Control of the caving position of overlying strata is crucial for successful retained entry from the principle of this non-pillar coal mining technology. To control the caving position, increase the height of the caving band, and cut off the stress transfer between the gob roof and the retained entry roof, He et al. invented the DRSB technique. 42,43 The DRSB technique refers to packaging several explosives into a bilateral cumulative device (BCD) with a symmetrical predetermined direction, as shown in Figure 8. When the explosive package blasts, the wall of the blast hole is uniformly compressed in the not designed energy cumulative direction, but the tensile force is concentrated in the designed direction, and the rock mass is cracked. The detonation energy is preferentially released from the accumulating hole, forming an energy flow under the guidance of BCD after the explosive is detonated, as shown in Figure 8A. The detonation energy first produces microcracks in the predetermined direction of the BCD, which provides pressure relief space for the detonation products because of the action of the BCD. When the compressive stress wave encounters the newly generated fracture surface, it will be reflected and become a tensile wave, which will cause tensile stress concentration in the direction vertical to the fracture surface, as shown in Figure 8B. The DRSB technique uses BCD to cumulative energy in a predetermined direction to split rock mass after considering the characteristics of rock tensile strength is much lower than compressive strength.
It can be concluded from the principle of BCD that the BCD has three mechanical functions: (1) generate uniform pressure in the not designed direction of BCD, and generate cumulative tension in the predetermined direction, as shown in Figure 8B; (2) the rock mass is partially compressed due to the energy accumulation of BCD, as shown in Figure 8C; On the xoz plane of the coordinate system, the rock mass between the holes is subjected to tensile stress after the surrounding rock is disturbed by the DRSB technique, in the axial direction of BCD, as shown in Figure 8D.
Based on the property of BCD, the DRSB technique is used to create a single crack surface in the retained entry gob-side roof before being disturbed by the influence of abutment pressure. A BCD with explosives is installed in a borehole designed at the retained entry gob-side roof, as shown in Figure 8A. When multiple boreholes are simultaneously explosive using DRSB, the cracks between each borehole will be connected and form a continuous single crack surface between the retained entry roof and the gob roof, as shown in Figure 9. Therefore, the DRSB technique can not only control the position of the caving gangue but also cut off the stress transfer from the gob roof to retained entry roof.

| Blocking-gangue support system
The BGSS is designed to prevent collapse gangue from falling into the retained entry and control the horizontal deformation of the gangue rib, as shown in Figure 10. There are two steps in the formation gangue rib process: the collapse process of overlying strata in the gob and the compaction process of gob materials. During the collapse of the overlying strata, the BGSS will be influenced by dynamic pressure shocks. During the gradual compaction process of the gob materials, the BGSS will be horizontally compressed. Therefore, BGSS is designed into three major parts: temporary gangue-blocking structure, sliding yield structure, and steel bar mesh. The temporary gangue-blocking structure is made of a steel plate. It is located behind the head of the coal drawing scraper conveyor and connected using bolts, as shown in Figure 10. It is designed to prevent the instantaneous impact of collapsed gangue and provide a safe space to install the sliding yield structure. Sliding yield structure mainly consisted of U-shaped steels, which are good at bending resistance and contractibility. Therefore, this sliding yield structure can adapt to the horizontal deformation of the gangue rib and the convergence of the roof and floor.
Before the overlying strata collapse, a temporary gangue-blocking structure is setup. Then the sliding yield structure is used to support the caving rock. First, the steel bar meshes are installed on the inner side of the temporary gangue-blocking structure to prevent the gangue from falling into the retained entry, as shown in Figure 11A. Then, the sliding yield structure is installed on the inner side of the steel  mesh to resist horizontal pressure and form the gangue rib. In addition, to prevent the single sliding yield structure from falling, a steel bar is welded to the top of the upper U-shaped steel. When installing the sliding yield structure, first insert the steel bars welded on the upper U-shaped steels into the blast holes, and then bury a part of the lower U-shaped steels into the floor. Moreover, it is necessary to use the connecting plate to connect all individual sliding yield structures into a whole structure to prevent the single sliding yield structure from falling. After the gob collapses, the temporary gangue-blocking structure blocks the gangue. As the mining panel advances forward, the temporary gangue-blocking structure and the coal drawing scraper conveyor move forward together, and the gangue rib is gradually formed and compacted, as shown in Figure 11C.

| Temporary support system
After the mining panel advanced forward, the overlying strata gradually collapsed and compacted, and a part of the retained entry behind the mining panel was in the dynamic pressure area. Therefore, TSS was designed to temporarily reinforce the support strength of the retained entry preventing the roadway from being affected by dynamic pressure and preventing the BGSS from falling before the gob material completely compacted. This TSS involves single hydraulic props, a top girder, and a bottom beam. It can provide temporary reinforced support in the retained entry and provide lateral support for the BGSS. The layout of the TSS is shown in Figure 11A. The principle of this TSS is as follows: Before the hydraulic support of the mining panel moves forward, install several rows of single hydraulic props along the strike direction of the retained entry, as shown in Figure 11A,B. The top end of each column of single hydraulic props is connected by the top girder, and the bottom end of each row of single hydraulic props is connected by the bottom beam. The top girder is to connect the single hydraulic props along the strike direction of the roadway and prevents the roof from being affected by the dynamic pressure. The bottom beam is a cross shape, which is used to prevent the single hydraulic props from sinking into the coal seam floor. Moreover, the bottom beam of the gob-side is close to the lower part of sliding yield structures, it can also prevent the BGSS from falling under horizontal pressure. Finally, the TSS can be withdrawn after the deformation of the retained entry is stabilized and the gob materials are compacted.

| The basic support of the roadway
The 150202 headgate is excavated along the coal roof, with a cross section of 5000 × 3000 mm. The supporting parameters for the roof of the designed retaining entry include the following: four threaded steel bolts of φ18 × 2200 mm are placed on the roof of the roadway with a row spacing of 780 mm × 800 mm; three regular anchor cables of φ18.9 × 6300 mm are placed on the roof of the roadway with a row spacing of 1560 mm × 1600 mm. The support parameters for the coal-mass rib including four fiberglass anchor bolts of φ22 × 1800 mm are placed on the coal-mass rib with the row spacing of 850 mm × 800 mm. The basic Support scheme of the 150202 headgate is designed as shown in Figure 12.

| The CRLD anchor cable reinforces support
During the process of using non-pillar coal mining technology to retain entry in the LTCC panel, the roadway experiences several deformation stages including the disturbance deformation caused by DRSB, initial coal mining, the roof caving, and gob materials compaction, and second coal mining. Therefore, it is necessary to use CRLD anchor cables to reinforce the support strength of retained entry before the extraction of the mining panel.
The parameters of CRLD anchor cable support for the retained entry include the following: a row of CRLD F I G U R E 12 Support scheme of 150202 headgate (unit in mm).
anchor cables with a dimension of 21.8 mm × 12,800 mm is placed on the gob-side roof with a row spacing of 800 mm. Four-row CRLD anchor cables with the dimension of 21.8 mm × 12,800 mm are placed on the roof with a row spacing of 3200 mm, as shown in Figure 13. The constant resistance value of the CRLD anchor cable is 350 kN, and the preload force should be greater than 25 t. Moreover, it is necessary to reinforce the support strength of the entry rib to prevent its falling because the large mining height will cause a more intense disturbance. Two rows of anchor cables with a dimension of 21.8 mm × 6300 mm are placed on the coal-mass rib with a row spacing of 1600 mm. The upper part of the coal-mass rib anchor cable is tilted 45°toward the entry roof with a distance to the entry of 800 mm. The upper part of the anchor cable is tilted 45°toward the entry roof with a distance to the entry of 800 mm. The lower part of the anchor cable is horizontally toward the coal-mass rib with a distance to the upper anchor cable of 1600 mm, as shown in Figures 13 and 14.

| The DRSB design
The DRSB technology is designed to cut off the stress transfer between the entry roof and the gob roof, increase the height of the caving band so that the collapsed gangue can fill the gob, and reduce the pressure of the retained entry roof. Based on EMT and Equations (5) and (6), the DRSB height H s shall be designed as follows 44 :  Figure 14. In addition, we also need to design the charge mass of a single hole and the distance of the adjacent hole. The final charge mass used in the single hole is The reinforced support strength design of retained entry (unit in mm).

F I G U R E 14
The profile of the directional roof splitting blasting design of retained entry. 5.6 kg after the field engineering test in the 150202 headgate. The installation method of emulsion explosives is 4 + 3 + 3 + 3 + 3 + 2 + 3 + 3 + 2 + 2 (number of emulsion explosive rolls, the weight of each roll of explosive is 200 g) in each BCD from the top to the bottom, as shown in Figure 15. The distance of the adjacent hole can be designed as follows 24 : where d is the distance of the adjacent hole, r b is the radius of the DRSB hole, μ is the Poisson's ratio of overlying rock mass, P b is the peak shock wave pressure on the blast hole wall, D 0 is the initial damage parameters of rock mass, and σ t is the tensile strength of rock. Therefore, r b = 24 mm, µ = 0.28, p = 7.8 MPa, P b = 2200 MPa, D 0 = 0.6, σ t = 16.5 MPa, substituting into Equation (10), the distance of adjacent hole d ≤ 542 mm. Therefore, the distance of the adjacent blast hole is designed to be 500 mm. The profile of the DRSB design is shown in Figure 14.

| The BGSS design
The overlying strata in the gob collapsed and gradually compacted after the mining panel advances. To block the collapsed gangue from entering the retained entry, the BGSS was designed to block the collapsed gangue and form the gangue rib. The BGSS includes three parts: temporary gangue-blocking structure, sliding yield structure, and steel bar mesh, as shown in Figure 16. The temporary gangue-blocking structure is bolted to the tail of the scraper conveyor behind the hydraulic support. It is made of a 10-mm thick steel plate with a size of 3000 mm × 2500 mm. The temporary gangueblocking structure can block the caving gangue shock disturbance and provide an area for installing the sliding yield structure and steel bar meshes behind the hydraulic support. The sliding yield structure is composed of two 2500-mm long U-shaped steels and the two U-shaped steels are overlapped and connected by two staples. During the construction of the BGSS, first, installt wo layers of steel bar meshes on the inner side of the temporary gangue-blocking structure (in the retained entry), and then install the sliding yield structure on the inner side of the steel bar mesh. The row spacing of the sliding yield structure is 500 mm. To prevent the sliding yielding structure from falling, the steel bars welded to the upper U-shaped steel are inserted into the blasting hole, then the lower U-shaped steel is buried 300 mm into the floor, and two rows of connecting plates are used to connect the adjacent sliding yielding structures.

| The TSS design
In the process of collapse and compaction of the overlying strata above the gob, the retained entry behind the mining panel will be disturbed by dynamic pressure. Therefore, it is F I G U R E 15 Schematic of charge structure (unit in mm). necessary to design the TSS in this disturbance area to reinforce the support strength of the retained entry and prevent the falling of the sliding yield structure. Based on previous engineering experience with approximate geological conditions, the length of the temporary support area behind the experimental 150202 mining panel was designed to be 300 m. The TSS includes three parts: single hydraulic props, top girder, and bottom beam, as shown in Figure 17. The TSS of the 150202 headgates behind the mining panel uses five rows of DW35-200 single hydraulic props, the DFB-3000 top girder, and the cross-shaped bottom beam. The length of the DFB-3000 top girder is 1000 mm. The bottom beam is welded into a cross shape by two I-beams, which can effectively prevent the single hydraulic prop from sinking into the coal seam floor. When installing the TSS, the top girder is installed on the top of the single hydraulic props along the strike direction of the retained entry and the bottom beam is installed at the bottom of the single hydraulic props along the dip direction. The gob-side bottom beam is close to the yield slip structure to prevent it from falling because the row spacing of the single hydraulic prop is 1000 mm, and the row spacing of the sliding yield structure is 500 mm, as shown in Figure 18.

| Analysis of practical application
The effect of non-pillar coal mining techniques applied to the LTCC panel is monitored from the pressure of single hydraulic props, CRLD anchor cable stress, roof separation layer, entry convergence, and gangue rib pressure. The arrangement of monitoring stations for hydraulic support, single hydraulic props pressure, CRLD anchor cable stress, roof separation layer, entrance convergence, and gangue rib pressure are shown in Figure 19. The deformation of the retained entry surrounding rock and the force on the support components are shown in Figures 20-23. A photo of retained entry formed by nonpillar coal mining technology in the LTCC panel is shown in Figure 25.
As shown in Figure 20, the stress on the two CRLD anchor cables is increasing abruptly as the mining panel advances forward in the abutment pressure range from −30 to −10 m. The stress of the CRLD anchor cable installed on the gob-side roof gradually increases and enters a state of constant resistance from −10 to 30 m behind the mining panel. The stress of the CRLD anchor cable installed on the coal-mass side roof gradually increased from −10 to 30 m but it did not reach the designed constant resistance of 350 kN. The stress of the two CRLD anchor cables suddenly decreases due to the collapse of the overlying strata above the gob from 30 to 80 m. At this stage, the stress fluctuation of the CRLD anchor cable installed on the gob-side roof is relatively large due to the disturbance of overlying strata collapse. Due to the supporting effect of the coal mass, the stress fluctuation of the CRLD anchor cable installed on the coal-mass side is relatively small at this stage. In the range from 80 to 180 m, the stress fluctuation of the two CRLD anchor cables decreases with the gob materials gradually compaction. At the position of 180 m behind the mining panel, the stress of the CRLD anchor cable begins to stabilize, and the stress of the CRLD anchor cable installed on the gob-side roof is greater than the coal-mass side roof.
As shown in Figure 21, in the range from 0 to 3 m behind the mining panel, the changes of stress in the gangue rib of retained entry are small. The stress in the gangue rib of retained entry is increasing as the mining panel advances forward from 3 to 70 m. The maximum pressure of the gangue rib is 1.8 MPa which reached 70 m behind the mining panel. This is because the overlying strata above the gob have completely collapsed in this position, but it has not completely compacted. The stress in the gangue rib suddenly decreases due to the gradual compaction of the gob materials from 70 to 180 m. At 180 m behind the mining panel, the stress in the gangue The arrangement of monitoring stations. rib of retained entry begins to stabilize at about 1.4 MPa. It can be seen from changes of stress in the gangue rib that the roof of the retained entry activity is most intense in the range from 0 to 80 m.
As shown in Figure 22, with the mining panel advancing forward, the stress curve of the single hydraulic prop gradually increases and is finally stabilized. The stress of a single hydraulic prop increased gradually due to the influence of the abutment press from −30 to 0 m ahead of the mining panel. At this stage, the stresses of the single hydraulic prop on the gob side and the coal-mass side gradually increased from 15 to 20.1 MPa and 19 MPa, respectively. As the mining panel advances, the stress of the single hydraulic prop increases rapidly from 0 to 30 m. However, in the range from 30 to 80 m, the growth rate of the single hydraulic prop stress decreased as the overlying strata collapsed. In the range from 80 to 180 m, the gob materials gradually compacted, and the hydraulic prop pressure increased slowly. At the position of 180 m behind the mining panel, stresses of the single hydraulic props on the gob side and the coal-mass side begin to stabilize at about 23 and 21.7 MPa, respectively. The highest stress of the single hydraulic props is 23 MPa, and the nominal design stress for the hydraulic prop is 25.5 MPa; therefore, the hydraulic single props can meet the requirements of a TSS. Figure 23 indicates that as the mining panel advances forward, the deformation of the retained entry surrounding rock gradually increases and finally stabilizes. The surrounding rock deformation of retained entry increased slowly due to the influence of abutment pressure from −30 to 0 m ahead of the mining panel. In the range from 0 to 30 m, the deformation rate of the retained entry surrounding rock increases. However, in the range from 30 to 80 m, the growth rate of the surrounding rock deformation decreased as the overlying strata collapsed in the gob. In the range from 80 to 190 m, the gob materials gradually compacted, and the deformation of retained entry is increasing slowly. Finally, at position 190 m, as the gob materials are completely compacted, and the deformation of retained entry is stabilized. The final surrounding rock deformation of the retained entry is as follows: floor displacement > coal-mass side displacement > gob-side displacement > roof displacement. The largest floor displacement is 314 mm, which meets the ventilation cross-section requirements. The floor displacement is mainly caused by the stress transfer from the overlying strata to the floor through the coal-mass rib.
The stress contours of the hydraulic supports using the LTCC and non-pillar coal mining technology are presented in Figure 24A,B. All the districts of the stress contour were divided into four zones, Zones I, II, III, and IV, according to the mining method and position of the hydraulic supports. The effect of pressure relief can be seen clearly when Zone II was compared with Zone IV along the strike direction. Zone II had a higherpressure area with a maximum pressure of 44.8 MPa, whereas Zone IV had a lower-pressure area with a maximum pressure value of 31.4 MPa. Compared with Zone II, the average pressure dropped to 27%, from 33.53 to 24.2 MPa in Zone IV. The maximum pressure of Zone I of the LTCC mining method was 43.5 MPa, and the change in pressure was small compared with Zone II along the dip direction. However, after using the nonpillar coal mining technology, the maximum and average pressures in Zone IV were 31.4 and 24.2 MPa, respectively. Compared with Zone III, the maximum and average pressures of Zone II decreased by 31% and 24%, respectively. These results demonstrate the pressure relief effect of the non-pillar coal mining technology. Figure 25 compares the overall effect of the roadway before and after using the non-pillar coal mining technology. It can be seen that the large deformation of the surrounding rock of the roadway is significantly reduced after using the non-pillar coal mining technology. This technology can not only control the large deformation of the roadway but also improve the recovery rate of coal resources in the coal mine. Therefore, this novel non-pillar coal mining technology is an effective method to realize safe and efficient mining in the LTCC panel. 5.6.2 | The safety and economy analysis 1. The safety aspect Compared with the conventional LTCC mining method, the non-pillar coal mining technology cuts off the stress transmission, eliminates the stress concentration, and finally, automatically retains the gob-side entry. Therefore, this non-pillar coal mining technology can effectively decrease the deformation of the retained entry and ensure safe production. Moreover, since the U-type ventilation system is changed into a Y-type ventilation system, it can effectively solve the problem of methane accumulating and exceeding the limit in the upper corner of the panel. In addition, this non-pillar coal mining technology can effectively reduce the loss of coal resources and can reduce the possibility of coal spontaneous combustion in the gob. The safety of this non-pillar coal mining technology is better than the traditional LTCC mining method.

The economic aspect
Compared with the conventional LTCC mining method, this non-pillar coal mining technology eliminates the coal pillars between two adjacent panels and reduces the 50% amount of roadway excavation work. Taking the 150202 LTCC panel as an example, the experimental panel can recover more coal resources of 145,656 tons and reduce the amount of roadway excavation works by 50% (1000 m) after using nonpillar coal mining technology. In summary, this novel non-pillar coal mining technology could increase the income is about 109 million RMB in the 150202 LTCC panel of the Sheilei coal mine. The economic benefits of the novel non-pillar coal mining technology are higher than the traditional LTCC mining method.

| CONCLUSIONS AND DISCUSSION
This research used a novel non-pillar coal mining technology for LTCC panels based on EMT. The four control techniques to control the surrounding rock of retained entry in the LTCC panel are introduced. Field tests show that this non-pillar coal mining technology can be used in LTCC panels. The main conclusions are as follows: 1. When the conventional LTCC mining method was adopted, the entry during the excavation process exhibit large deformation. Furthermore, this method needs a large amount of roadway excavation, which will cause the excavation relay out of balance. Moreover, the large mining height of the LTCC mining panel will cause a more serious loss of coal pillar resources. To address these problems in traditional LTCC mining panels, this research proposes a non-pillar coal mining technology for LTCC panels based on EMT. 2. When the non-pillar mining technology was applied in LTCC panels: First, the entry is excavated along the roof of the coal seam, to facilitate support retained entry. Moreover, four key control techniques were applied to control the surrounding rock of retained entry: CRLD anchor cable, DRSB, BGSS, and TSS. The CRLD anchor cable could accommodate the large deformation and play a significant effect. The DRSB technique was adopted to increase the height of the caving band, control the caving position of the overlying strata above the gob, and cut off the stress transfer between the gob roof and retained entry roof. When the mining panel moves forward, the overlying strata in the gob will collapse along the splitting line. The application of BGSS could prevent collapsed gangue from falling into the retained entry and control the deformation of the gangue rib. TSS was being used to temporarily reinforce the support strength of the retained entry and prevent the BGSS from falling. Finally, in the position of 180-190 m behind the mining panel, the field monitor indicates the gob materials are completely compacted, and the deformation of the retained entry is stabilized. 3. The experiment of non-pillar coal mining technology in the LTCC panel shows that this technology could reduce the amount of mining entry excavation work by 50%, and eliminate the coal pillars and increase the coal resource recovery rate of the mining area by up to 13%. Furthermore, the gob materials will become a natural support body after compaction and stability, which will keep the roadway in a low-stress zone. This technology can effectively control the large deformation of the mining roadway. Therefore, this non-pillar coal mining technology should be promoted in thick coal seams, especially in the LTCC mining panel.

| SUMMARY
At present, with the growth of energy demand and the improvement of coal production, more and more coal mines are encountering the problem of excavating relays out of balance and the low recovery rate of coal resources. To solve this problem, a novel non-pillar coal mining technology has been proposed. This technology proposes to automatically form a gob-side entry by utilizing the ground pressure. Compared with the GER method, this technology eliminates to reservation of packing walls or coal pillars. Moreover, due to the bulk factor property of gob materials, the collapse gangue could fill the gob and support the overlying strata, so the retained entry is kept in the pressure relief area. Therefore, this technology has completely changed the stress environment of the surrounding rock of the retained entry and reduced the large deformation of the retained entry. At present, this method has been matured in the full-height mining panel whose mining height is less than 4.5 m (see the introduction part of this article for the application case). However, as the increase of mining height, the mine pressure appearing on the working face of thick coal seams and extra thick coal seams is more intense. In addition, there are some differences between the LTCC mining method and the full seam mining method. Therefore, it is necessary to study the non-pillar coal mining technology in LTCC panels. Compared with the full seam mining method, the non-pillar coal mining technology in LTCC panels has the following differences: (1) To reduce the horizontal stress of BGSS, this manuscript proposes that the roadway is excavated along the roof of the coal seam; (2) due to the increase in the mining height, the mine pressure appearing on the working face is more intense. Therefore, the authors propose using the anchor cable to reinforce the support strength of the coal-mass rib before retaining the roadway; (3) Since scraper conveyors are installed at the front and rear of the hydraulic support of the LTCC panel, the BGSS needs to be constructed behind the working face. Therefore, the temporary gangue-blocking structure is specially designed to protect the construction of the BGSS in this paper; (4) to prevent the single hydraulic props from sinking into the coal seam floor, the cross-shaped bottom beam and single hydraulic prop are specially designed for temporary support in the retained entry. Finally, the field engineering application verified the feasibility of the support scheme of the non-pillar coal mining technology in LTCC panels.
However, this study is only based on the non-pillar coal mining technology in LTCC panels of the Shenlei Coal Mine. However the geology and mining conditions of each coal mine are different, so it is necessary to study the applicability of the proposed method in other coal mines. Furthermore, it is necessary to study the design methods of CRLD anchor cable support parameters, DRSB height, and TSS parameters of the non-pillar coal mining technology in LTCC panels.