《1. Introduction》

1. Introduction

China continues to dominate the global market for coal, and the outlook for the next 20 years indicates that China will remain the world’s largest consumer of coal [1]. Based on insights into China’s energy consumption, coal consumption will still account for over 50% of primary energy by 2030 [2,3]. Meanwhile, coal demand within India and other emerging Asian economies will increase, since coal will be used to meet robust growth in power demand as these economies grow and their prosperity increases [4]. China is the biggest coal-mining country and a global pioneer in mining technology, and its longwall mining technology development is greatly improving coal production [5]. However, China faces a significant problem of coal resource waste during exploitation. The mine average recovery rate is only about 50%, as many coal pillars acting as shields are left underground and cannot be reclaimed [6,7]. In Fig. 1(a) [8], a long coal pillar is set between mining panels I and II, and the pillar width would generally increase as mining activity continues to a deeper level [9]. Furthermore, the width design of the entry pillar is a complicated issue that has come under continual study around the world [10–16].

Therefore, this study presents a non-pillar longwall mining technology that does not require the entry pillar setting. As shown in Fig. 1(b) [8], the II head entry will be automatically formed from the I tail entry during panel I advances. Using the technology presented here, this automatically formed entry can be safely and stably retained for subsequent panel mining. In this way, no coal pillars are set between the mining panels, and coal resources can be extracted to the greatest extent, allowing engineers to avoid the protective coal pillar retention problem. Furthermore, the head entry of the next mining panel is automatically formed during the mining of the previous panel. The preparation entry drivage work is cut in half compared with the conventional mining method, significantly decreasing the workload and thus reducing the potential drivage hazard [17–20].

《Fig. 1》

Fig. 1. Schematic diagram of the conventional and non-pillar mining approaches. (a) Conventional entry layout; (b) entry layout for non-pillar mining. Reproduced from Ref. [8] with permission of Elsevier Ltd., ©2019.

In recent years, industrial experiments on non-pillar mining at pilot sites have been successfully conducted, led by the State Key Laboratory for Geomechanics and Deep Underground Engineering, Beijing, China. Guo et al. [21] examined the feasibility of this technology in thin coal-seam mining, and He et al. [22] studied the adaptability of this technology for medium-thickness coal-seam mining. He et al. [23] conducted an industrial test of this technology in thick coal-seam mining with a fast mining rate and achieved a satisfying effect. He et al. [24] also successfully applied this technology to deep coal mining.

In this paper, we systematically summarize this non-pillar mining technology. The principles behind the technology are analyzed first. The key techniques involved in the technology are then introduced; their design methods and application effects are studied separately in combination with their field application. Finally, the retained entry stability and engineering effect are discussed.

《2. Principles of non-pillar mining》

2. Principles of non-pillar mining

The automatically retained entry is utilized by the next panel, and no coal pillar is left during the mining. Due to the strata movement, the retained entry faces intense mining pressure. In this technology, we use three key techniques to ensure entry stability: constant-resistance large deformation (CRLD) anchors, directional presplitting blasting (DPB), and a blocking-gangue support system (BGSS). As shown in Fig. 2, CRLD anchors are first applied to support the entry roof. DPB is then applied in the roof on the mining side. A smooth fracture face (presplitting plane) is formed under the DPB effect. As the coal seam is mined out, the roof within the DPB range caves under the mine pressure. The caved rock mass expands because of the broken-expansion nature of the rock. The expanded rock material compensates for the coalextraction space. Therefore, the upper roof movement is restricted. Meanwhile, the caved rock mass in the gob becomes the natural rib of the entry. The BGSS is set at the gob side in the entry to mold the integrated rib. This entry is automatically formed and safely retained to serve the next panel. This method allows the maximum extraction of coal resources in the mining area, and reduces the entry excavation by half in the subsequent mining. In the practice of longwall mining, entries are excavated in advance to prepare the mining face. Therefore, CRLD support and DPB can well be preimplemented after the entry excavation. When the mining starts, the BGSS is installed behind the mining face and is implemented simultaneously with the advance of the mining face. Therefore, the implementation of these three techniques can be prepared well and the techniques do not interfere with each other, ensuring mining efficiency.

《Fig. 2》

Fig. 2. Schematic diagram of non-pillar coal mining with an automatically formed gob-side entry. (a) Three-dimensional (3D) view of the retained entry; (b) stratigraphic model of non-pillar mining with a retained entry.

《3. Key techniques》

3. Key techniques

《3.1. CRLD support》

3.1. CRLD support

The CRLD anchor cable was developed and patented by Manchao He and his research team [25,26]. In actual engineering application, CRLD anchors can accommodate large deformation of the adjoining rock mass at a great depth in response to external force. The anchor consists of two parts: the constant-resistance body and the bolt shank. As shown in Fig. 3 [25], the constantresistance body is composed of a cone unit and a sleeve. The sleeve acts as a slide track for the cone unit. The shank is anchored by grouting in the depth of the rock mass—that is, in the fixed stable region. On the surface of the anchored mass, a combination of a face pallet and tightening nut is used to fix the free end. When the rock mass deforms under external disturbances, the constant-resistance body generates an internal slide, and the sliding distance depends on the free length of the CRLD anchor. At present, the free length is from 300 to 2000 mm of the different CRLD specifications [27]. Three stages of the sliding movement are illustrated in Fig. 3 [25]: the elastic deformation stage (Fig. 3(a)), constant-resistance deformation stage (Fig. 3(b)), and ultimate deformation stage (Fig. 3(c)). At the elastic deformation stage, the axial force caused by the rock deformation is less than the constant resistance of the CRLD anchor, which is not enough to activate the cone unit sliding in the sleeve. The elastic deformation is tiny and occurs within the constant-resistance body and bolt shank themselves; the bolt does not elongate substantially. As the axial force increases to the constant force, the CRLD anchor enters the constant-resistance deformation stage. The CRLD anchor maintains high constant resistance during bolt elongation (i.e., the sliding movement of the cone unit). This resistance is predefined by the function of the cone unit and sleeve. At present, the successfully tested resistance is up to 850 kN [27]. Therefore, the CRLD anchor absorbs a massive amount of energy to resist the consistent deformation and failure of the country rock mass in the constantresistance deformation stage. The elongation will eventually stop after the energy is fully released; at that moment, the external force will be smaller than the constant resistance. The rock mass within the anchored range will achieve a new stable state after the strong disturbance.

《Fig. 3》

Fig. 3. Working principle of the CRLD anchor. (a) Elastic deformation stage; (b) constant-resistance deformation stage; (c) ultimate deformation stage. Reproduced from Ref. [25] with permission of Elsevier Ltd., ©2014.

An analytical load-elongation relation was established for the CRLD anchor according to its constitutive relation [25]. Fig. 4 [25] shows several typical cycles of the analytical load-elongation curve for the CRLD anchor based on the calculation of a 16 t CRLD anchor. In the initial stage (elastic deformation stage), the resistance elastically increases with a tiny displacement of less than 20 mm. The curve of the elastic deformation stage is in accordance with Hooke’s law, P = kx, where P is static tensile load, k is the stiffness of the bolt shank, and x is the displacement or elongation. When the increased force achieves the predesigned constant resistance, the curve oscillates periodically in the constant-resistance zone with the continuously increasing displacement. The calculated maximum and minimum forces are 180 and 140 kN, respectively. These two limit values remain stable while the CRLD anchor is elongating. Related laboratory tests were conducted and developed to observe the CRLD performance, and the test results verified the ability of the anchor to accommodate a large deformation with a high constant resistance [28].

《Fig. 4》

Fig. 4. Analytical load-elongation curve of the CRLD anchor. The anchor’s size is given in the sketch (unit: mm). ω: Undamped natural frequency; f: static frictional coefficient; fd: dynamic frictional coefficient; f' : equivalent frictional coefficient; k: shank stiffness; Is: sleeve elastic constant; Ic: cone geometrical constant; x0: elastic displacement; : cycled displacement. Reproduced from Ref. [25] with permission of Elsevier Ltd., ©2014.

《3.2. Directional presplitting blasting》

3.2. Directional presplitting blasting

The DPB technique applied in the new non-pillar mining approach is based on the bilateral cumulative explosion technology presented and developed by Manchao He and his research team [29,30]. This technology is aimed at directionally blasting a material that has a high compression resistance and low tension resistance. This technology makes use of a bilateral energygathering device. The explosive blasts in this device and the blasting energy are converted into point-strip energy flow via energygathering holes. As shown in Figs. 5(a) and (b) [31], the ejected point-strip energy flow applies the cumulative tension on the local area of the borehole (i.e., the area of the energy-gathering holes) while the remaining area of the borehole is uniformly compressed due to the protective function of the energy-gathering device. Therefore, a directional crack can be developed in material that is good at resisting compression but fails under tension. The rock itself possesses this mechanical property. A blasting test was conducted in the rock mass, and the application effect is illustrated in Fig. 5(c) [31]. A line of boreholes using the bilateral cumulative explosion technology were blasted together in the rock mass. A directional crack connected these boreholes along the energygathering direction, and no other visible cracks were generated in other directions.

《Fig. 5》

Fig. 5. Mechanical model of the directional blasting and its application effect. (a) Cumulative blasting diorama; (b) cumulative blasting effect in the view of the x–z plane; (c) multi-hole blasting effect in rock mass. Reproduced from Ref. [31] with permission of Elsevier Ltd., ©2020.

DPB is used to generate a smooth structural surface between the retained entry roof and the gob roof before the mining activity arrives. As shown in Fig. 6 [8], bilateral energy-gathering devices with the explosive are installed into boreholes that are designed in the retained entry roof on the mining side (i.e., the gob side after mining, as shown in Fig. 1 [8]). Rows of the energy-gathering holes are aligned along the roadway strike direction. By setting a specific interval among these devices, a presplitting plane is generated in the energy-gathering direction by means of the bilateral cumulative explosion technology. DPB realizes the separation between the retained entry roof and the gob roof, which artificially controls the caving position of the gangue on the entry side. This makes it possible for the caved gob roof to turn into the rib of the retained entry. Moreover, as a refined blasting technique, DPB will not damage the original roof integrity of the retained entry.

《Fig. 6》

Fig. 6. DPB application in a retained entry roof. Reproduced from Ref. [8] with permission of Elsevier Ltd., ©2019.

《3.3. Blocking-gangue support system》

3.3. Blocking-gangue support system

To integrate the caved rock material on the gob side into an effective entry rib, the gob-side support technique was studied. In terms of the space–time relation, the dynamic course of the caved material exists in two forms: the caving process and the compacting process. The falling rock material first causes an instantaneous impact on the gob-side support in the caving process and then causes a lateral extrusion to the gob-side support in the compacting process. The BGSS is accordingly designed with three major parts: an anti-impact self-advancing structure, a sliding-yield structure, and an auxiliary supporting structure. The structure layout of the BGSS is shown in Fig. 7. The anti-impact self-advancing structure is located right behind the face-end support and is connected to it; a metal mesh is set to segregate the gangue; and the sliding-yield structure and auxiliary supporting structure are spaced reciprocally outside the metal mesh. First, the anti-impact self-advancing structure in the rock caving area converts the local impact into integral load-bearing by increasing the force area with the gangue and the contact area with other structures; this reduces the impact on the individual blockinggangue structure. Moreover, this structure realizes selfadvancement by connecting with the face-end support for timely resistance of the instantaneous impact in the caving process. The sliding-yield structure is composed of overlapped U steel, which possesses excellent resistance against bending. The sliding-yield structure can also slip appropriately to accommodate vertical deformation due to roof pressure during the compacting process. Adjusting the torque of the clips can strengthen the structure axial bearing capacity. These performances ensure the integrity and reusability of the sliding-yield structure. The auxiliary supporting structure is used to resist the roof pressure on the presplitting side, thus reducing the axial load on the sliding-yield structure. An excessive axial load would cause the structure to bend locally and would influence the structural resistance to lateral deformation. Therefore, the setup of the auxiliary supporting structure potentially maximizes the resistance to the lateral deformation of the sliding-yield structure in this circumstance. In addition, the substantial contact between the sliding-yield structure and the entry roof and floor generates resistance friction to control the gangue lateral deformation cooperatively. According to different geological mining conditions, we designed and adopted a matched auxiliary supporting structure. The hydraulic prop is applicable to coal-seam mining of thin and medium thickness, while the unit support is applicable to thick coal-seam mining, whose rock pressure phenomenon is more violent. The constructions of the BGSS structures are shown in Fig. 8.

《Fig. 7》

Fig. 7. Layout of the BGSS structures.

《Fig. 8》

Fig. 8. Classification and construction of the BGSS structures.

《4. Field application》

4. Field application

《4.1. Site conditions》

4.1. Site conditions

The Baoshan coalmine, located in Inner Mongolia, China, was selected for the field application of this technology. As shown in Fig. 9(a), panel 6301 became the gob after mining, and the coal pillar was left between the 6301 and 6302 mining panels. The proposed non-pillar mining technology was adopted during the mining of panel 6302, so the 6302 tail entry was retained automatically as the 6303 head entry for the next panel; thus no coal pillar was left there. The roof lithology 10 m above the 6302 tail entry and the lithology of the entry were investigated, as shown in Fig. 9(b). The entry was 2.45 m in height and was a half-coal and half-rock tunnel excavated along the top of the coal seam; its average buried depth was 60 m. The mean thickness and inclination of the mined coal seam were 1.56 m and 2°, respectively; thus, it was a medium-thickness and near-horizontal coal seam. The immediate entry roof was fine sandstone; it was thus a hard rock roof. The upper roof and the floor of the entry were sandy mudstone with medium strength. The mining panel was 200 m wide along the dip direction and 890 m long along the strike direction, so the retained length of the 6302 tail entry was 890 m.

《Fig. 9》

Fig. 9. Mining panel layout and geological conditions. (a) Layouts of the 6302 mining panel and its adjacent panels; (b) roof lithology for the retained entry.

《4.2. Field methods and designs》

4.2. Field methods and designs

4.2.1. CRLD support design

During the course of the application of this technology, the structural conditions of the retained entry roof varied considerably as the mining panel advanced. The initial roof state, which was free from any mining disturbance, was the most stable. According to the principles of this technology (Section 2), DPB was applied before mining, and divided the entry roof and the gob roof. As the coal was extracted, the retained entry roof lost the mining side support and hung temporarily, due to the spatiotemporal behavior of the gob roof caving. After the broken rock expanded sufficiently, the entry roof touched the gangue and acquired natural support. Therefore, the entry roof in the gob roof-caving area was the least stable over the entire process. To facilitate analysis of the mechanical roof states, we established rock beam models without considering the support conditions, as shown in Fig. 10. In its most stable state, the rock beam is regarded as a fixed beam; in its least stable state, it is regarded as a cantilever beam. Taking the rock roof dimensions as long, h deep, and 1 m wide, the elasticity solutions of the axis deflection (v) for both models were respectively calculated as follows:

where q1 and q2 are the uniform loads on the fixed rock beam and the cantilever beam, respectively; E is the elastic modulus of the rock mass; and μ is Poisson’s ratio of the roof rock mass.

《Fig. 10》

Fig. 10. Mechanical state evolution of the retained entry roof. (a) Initial state before mining; (b) working face roof-caving state after mining. q1 and q2: the uniform loads on the fixed rock beam and the cantilever beam, respectively; E: the elastic modulus of the rock mass; μ: Poisson’s ratio of the roof rock mass; h: the height of the rock beam; v: the axis deflection of the rock beam.

In the initial state before mining, maximum deformation of the entry roof occurred in the middle. Shortly after mining, the maximum deformation was transferred to the edge on the roof splitting side; the configurational freedom of the rock roof increased.

Based on the above analysis, CRLD anchors were first installed to protect the entry roof. A row of CRLD anchors at intervals of 1 m was installed on the roof splitting side, and the anchors were connected with W-steel belts for cooperative control. In addition, a row of CRLD anchors at intervals of 3 m was installed on the middle of the roof to reinforce the original support. As shown in Fig. 11, two rows of CRLD anchors and W-steel belts were added based on the original support. Since the burial depth of the entry was comparatively shallow, the mine pressure was not great. The specifications of the CRLD anchor were 300 mm free length and 25 t constant resistance.

《Fig. 11》

Fig. 11. Retained entry roof support design. (a) Unfolded drawing of the roof support (unit: mm); (b) field scene. W-steel specification: 2400 mm × 280 mm × 4 mm. Φ is the diameter of the steel strand of bolt or anchor.

4.2.2. DPB design

DPB in the field is designed to separate the roof and make the caved gob roof compensate for the mining void. Therefore, the DPB height and angle in the roof should be specifically designed based on the conditions of the site. First, the DPB height H should satisfy the following:

where m is the mining height, kb is the bulking factor of the rock roof, and θ is the DPB angle.

The DPB angle is set to make the rock roof within the DPB range collapse effectively and rapidly, so that the caved rock material becomes the entry rib and expands quickly to support the cantilever rock beam. The caved rock material within the DPB range is located on the lower roof and fails in a sliding manner. According to the instability principle of a voussoir beam [32], the sliding instability of the interacted rock beams occurs when

where φf is the rock friction angle, h0 is the rock block height, is the rotational subsidence of the rock block, and L is the rock block length.

Substituting the related geological parameters of the field surrounding rock into Eq. (4), where φf = 30°, h0 = 3.78 m, = 1.6 m and L = 15.5 m, we obtained θ ≥14.28°. The greater the angle is, the higher the DPB length will need to be. The practical DPB angle was determined to be 15°. Therefore, H ≥4.73 m was obtained according to Eq. (3), where m = 1.6 m and kb = 1.35. Considering the roof strata relation from Fig. 9 and the operability, the practical DPB height was determined to be 5 m. After determining the DPB borehole length, the charging parameters and the hole distance needed to be designed. The charging parameters are generally determined by site tests for the optimal charge quantity. The ultimate charge structure was tested to be a "3 + 2” pattern, in which emulsion explosives with a unit length of 300 mm were arranged in a decoupling air-spacing way; the detonation mode was serial blasting. The distance between boreholes can be derived by the following [33]:

where d is the hole distance; rb is the hole radius; ρ0 is the explosive density; Dj is the detonation velocity; λ is the coefficient of the side pressure; n is the enhancement coefficient of detonation products; ξ is the energy-focusing blasting coefficient; D is the damage variable of rock mass; γb is a constant related to explosive property and charging density; σt is the rock static tensile strength; γ is the rock density; Hb is the buried depth of the rock; le is the summation length of explosives; lb is the length of charging segments; c is the ratio of the diameters of the borehole and the explosive.

Fine sandstone accounted for the majority of the rock within the 5 m DPB height (Fig. 9), and the tensile strength of the sandstone was greater than that of the sandy mudstone (the remainder of the rock mass within the DPB range). Therefore, according to the physical–mechanical properties of the sandstone and the used explosive specification, d ≤ 518.77 mm was calculated by substituting the following parameters into Eq. (5): rb = 24 mm, ρ0 = 1200 kg·m-3 , Dj = 3600 m·s-3 , λ = 2.6, n = 10, n = 2, ξ = 0.7, γb = 3, σt = 2.6 MPa, γ = 25 kN·m-3 , Hb = 60 m, le = 1.5 m, lb = 3.0 m, and c = 0.75.

The DPB design overview is shown in Fig. 12. Note that the designed height of the CRLD anchor should be greater than the DPB height, and the difference is generally 2–3 m, which allows the fixed length (Fig. 3 [25]) to be free from the blasting influence. In addition, the CRLD anchor can firmly hang the immediate roof of the retained entry on the thick and strong upper rock formation.

《Fig. 12》

Fig. 12. Section diagram of the DPB retained entry and DPB design.

4.2.3. BGSS design

The anti-impact self-advancing structure connected with the face-end support was made of deformable steel with a length of 6 m and a height of 1.5 m. This structure was behind the metal mesh and prevented the metal mesh from being damaged by the caving gangue. The 100 mm × 100 mm metal mesh was used to integrate the gangue wall and prevent the small gangue from thrusting into the entry. The sliding-yield structure was made of overlapped double U steel, which nicely adapted to the vertical deformation and withstood the horizontal deformation. Therefore, the double U steel used for long-term support could be used again to serve the next entry after the retained entry was abandoned. The auxiliary supporting structure comprised a hydraulic prop. As the temporary support, the hydraulic props were set in the dynamic pressurebearing zone, which stretched for an empirical length of 150–200 m behind the working face. Thus, the props were used and reused as the working face advanced. The hydraulic prop and double U steel were placed in a staggered arrangement at intervals of 500 mm. The BGSS design for the field is shown in Fig. 13.

《Fig. 13》

Fig. 13. BGSS design in the field.

《4.3. Field monitoring》

4.3. Field monitoring

4.3.1. CRLD support effect

As discussed in Section 4.2.1, the entry roof would deflect the most on the roof-splitting side. Choosing to place the CRLD anchor on the splitting side, we monitored the CRLD anchor stress and its retraction value during the face mining. The stress was monitored in real time by a YAD-200 vibrating-string cable dynamometer that is manufactured by Shandong University of Technology Zhongtian Safety Control Technology Co., Ltd., China, and the retraction value was measured continuously with a Vernier caliper. The dynamometer outputted the anchor loads Ri based on the calculation formula without regard to temperature change:

where G is the apparatus coefficient, f0 is the initial frequency modulus of the vibrating string, and fi is the real-time recorded frequency modulus of the vibrating string.

The comprehensive performance of the CRLD anchor is shown in Fig. 14. A preloading force no less than 250 kN was first applied to the CRLD anchor during the installation. At the initial stage of the monitoring area (–20 to 3 m), when the position of the CRLD anchor ranged from 20 m ahead of the mining face to about 3 m behind the mining face, the CRLD anchor was in the constantresistance state and no retraction occurred; the output stress fluctuated smoothly. After that, the stress vibrated dramatically, and the vibration amplitude decreased as the mining face advanced. Meanwhile, the retraction value increased rapidly; the constant-resistance body slid accordingly. When the mining face was about 52 m ahead of the CRLD anchor, the anchor stress leveled off again; the corresponding retraction value no longer increased significantly. The recorded ultimate retraction value stabilized at 28 mm. As seen from the above phenomena, the most intense activity period of the retained entry roof occurred within 60 m behind the working face. The CRLD anchor was well adapted to the large deformation of the entry roof and displayed superior energy absorption while resisting roof sagging.

《Fig. 14》

Fig. 14. CRLD anchor support performance.

4.3.2. DPB effect

The CXK-6 borehole imager, which is manufactured by Wuhan Conourish Coalmine Safety Technology Co., Ltd., China was used to observe the formation effect of blasting cracks to optimize the charging parameters and supervise the blasting quality. As shown in Fig. 15(a), two clear directional cracks were generated during the charging passage (for a borehole depth of 2–5 m) by using the '3 + 2” pattern described in Section 4.2.2. In addition, the DPB boreholes should be subjected to a spot check of the crack ratio (the crack length divides the borehole length), which is expected to be higher than 60%. The crack ratio in the field was 74% under a random check every 200 boreholes (i.e., every 100 m along the entry strike direction). As shown in Figs. 15(b) and (c), the DPB effect could be observed from the automatically formed entry rib. DPB separated the roof along the borehole line, and the half-hole on the gob roof side would cave in after mining. We captured the rock fracture plane with the blasting hole. The half-hole was left in the gob area; no other apparent cracks were formed on the borehole surface, and the fracture plane was smooth. These occurrences demonstrated the expected DPB application effect.

《Fig. 15》

Fig. 15. Directional blasting effect. (a) Borehole imaging; (b) half-hole in the caved zone; (c) enlarged view of the half-hole.

4.3.3. The BGSS effect

To ascertain the change in lateral gangue pressure applied on the BGSS as the working face advanced, we set the pressure gauge behind the double U steel to record the pressure change, as shown in Fig. 16. In the rising stage, the pressure appeared after the working face had advanced by 4.2 m. The pressure then slightly increased from 0.33 to 0.38 MPa between the advanced distances of 4.2–8.2 m, while the pressure climbed fast from the distance of 8.2 m; the maximum pressure was 1.63 MPa at 45.2 m. These results indicated that the hard roof caving had an obvious space lag when the working face was pushed away. The anti-impact structure worked and decomposed the impact force from the gob roof first caving, which corresponded to the early slow increase period. After this period, the anti-impact structure moved forward and the upper roof collapsed in layers, which led to a rapid pressure increase. In the falling stage, the pressure declined after the impact motion of the upper main roof, and then gradually became stable at about 1.22 MPa at around 96 m. At this point, the compacting gangue was basically stable.

《Fig. 16》

Fig. 16. Gangue pressure monitoring curve.

The retained entry segment behind the working face was divided into two parts: the dynamic pressure-bearing zone and the entry stabilization zone. In the dynamic pressure-bearing zone, the double U steel and the hydraulic prop were alternately spaced, as shown in Fig. 17(a). The caved rock material was blocked to form the integrated entry rib by the BGSS. As the working face advanced, the segment that was previously in the dynamic pressure-bearing zone would enter the entry stabilization zone, and the props therein would be retracted to support the next dynamic pressure-bearing zone. As shown in Fig. 17(b), the retained entry in the stabilization zone was already steady, and the double U steel did not have apparent lateral deformation after the prop retraction. The final automatically formed entry rib met the production and safety requirements.

《Fig. 17》

Fig. 17. Field application effect of the BGSS. (a) Before the prop retraction; (b) after the prop retraction.

4.3.4. Mining pressure

By processing the pressure data of the working face supports, we obtained the three-dimensional (3D) distribution of the working face mining pressure, as shown in Fig. 18. The mining pressure was minimal on the retained entry side. The pressure increased along the dip direction of the working face, and the maximum pressure was on the non-retained entry side. These results demonstrated that, due to the DPB design and construction, the caved rock material fully expanded to become a natural supporting body and restricted the increase of mining pressure. This pressure reduction effect decreased along with the increasing distance away from the DPB position. Therefore, the maximum pressure was located on the non-retained entry side, where the entry was under the conventional mining way. Thus, the retained entry faced less mining pressure in this non-pillar coal-mining technology than it would have under traditional mining technology, which was very beneficial to the entry stability.

《Fig. 18》

Fig. 18. 3D nephogram of the mining pressure.

4.3.5. Retained entry stability

The roof-to-floor displacement of the retained entry is a direct index reflecting the entry stabilization. We set displacementmonitoring points in the retained entry in order to observe the roof-to-floor convergence change. As shown in Fig. 19, the displacement gradually went from rising to stable as the working face advanced away from the measuring point. The displacement rising phase was defined as the dynamic pressure-bearing zone, and the phase when the displacement tended to be stable was defined as the entry stabilization zone. In the dynamic pressure-bearing zone, the retained entry was close behind the mining face. Because of the mining disturbance and the gob upper roof movement, the retained entry was subjected to converging forces and generated convergence. Therefore, we set the temporary support (hydraulic props matching up with the lace girder) in this zone to reduce the entry displacement and promote a transition to the stable stage. The monitoring result showed that the stabilization distance was around 148 m on site. When the entry was entering the stable stage, the props could be withdrawn and utilized to support the newly formed dynamic pressure-bearing zone. In the following field operation, the prop retraction distance at the site was set at 160 m to be on the safe side. With such a retraction circulation, the whole 890 m entry in the field was retained successfully. The roof-to-floor displacement stabilized at 212 mm, which met the entry retaining and mining production requirements.

《Fig. 19》

Fig. 19. Roof-to-floor displacement and retained entry effect.

《4.4. Problems of retained entry》

4.4. Problems of retained entry

During the field application of the technology for a mediumthickness coal seam with a hard roof, a few sections along the retained entry showed a coal rib spalling and a rock arch hanging in the gangue rib, as shown in Fig. 20. First, the coal rib spalling of the retained entry increased the entry span, which led to an increase of the unsupported area of the entry roof and a decrease of the roof safety factor due to the long service life of the entry. From the investigation and analysis, we considered that, in this segment, the poor connectivity of the directional cracks caused the entry rib to bear more load from the movement of the cantilever rock beam. When the gob roof caved along the poor presplitting plane, the entry roof had to overcome the cohesive force in the uncut positions. Because the roof rock was hard, the entry roof deflected more and thus squeezed the coal rib, causing the spalling. Another problem of the hanging rock arch was a potential threat to entry safety. We believed that there were two reasons behind that: First, DPB did not cut apart the roof effectively; and second, the fractured lumpiness of the rock roof was large and the rotary space was small due to the site conditions of a hard roof and a medium-thickness coal seam. When the inside edge of the big rock touched the floor after its small rotation, this rock was balanced by another edge of the friction on the presplitting plane. This balance would be easily broken when the next mining face approached. Once the big rock fell, the regional gob-side support might be destroyed, which would be a threat to the moving workers. However, adjusting the measures to the site conditions solved these problems: The connectivity rate between the boreholes was elevated, the entry coal rib was supported, and the loose blasting boreholes next to the DPB boreholes were increased in order to decrease the rock lumpiness.

《Fig. 20》

Fig. 20. Problems during the entry retaining. (a) Coal rib spalling; (b) rock arch hanging in the gangue rib.

《5. Discussion》

5. Discussion

Pillar-less longwall mining first began to appear in the 1950s. The conventional method is to build the gob-side pack; this method has been tested and employed under many geological conditions for different mining depths [34,35], coal-seam thickness [36,37], and roof lithology [38–40]. As a pioneering approach in pillar-less longwall mining, this conventional approach has been developed considerably over the past decades. However, some inherent disadvantages have emerged. Long-term roof movement disturbances make the retained entry difficult to maintain under the conventional approach [41]. Furthermore, the application conditions are limited; for example, this method has bad adaptability in the case of a hard roof [42]. These problems are becoming intractable as mining depths increase [43]. Therefore, it is complicated to apply the conventional approach in practice [44]. In addition, building the gob-side pack by means of construction or material filling has a high cost in both workforce and material resources, and introduces potential delay that reduces the efficiency of longwall mining.

This study presents the innovative technology of non-pillar longwall mining with an automatically formed entry, which uses the self-bearing ability of the gangue to relieve the mining pressure and form a natural entry rib. Based on the proposed design principles, three key techniques (CRLD, DPB, and BGSS) cooperate to retain the gob-side entry efficiently and safely. Some problems may be encountered during actual application due to different geological mining conditions. The ideal application effect can be achieved by adjusting the measures to the specific site conditions.

Coal distribution and mining conditions are complicated [45]. For potentially disastrous mines (e.g., those with risk of rockburst, coal and gas outburst, and mine water disaster), this innovative technology has good application potential, although it still needs to be further studied and tested in the field.

This innovative non-pillar coal-mining technology takes the proposed principles as the framework and requires the adjustment of some measures to different site conditions to ensure the quality and safety of the retained entry. This technology has good engineering applicability and promotion value.

《6. Conclusions》

6. Conclusions

A non-pillar coal-mining technology with an automatically formed entry was studied, which reduces the waste of coal resources and entry excavation. Three key techniques involved in the technology were introduced, namely: CRLD support, DPB, and BGSS. These three designs and their effects were investigated by means of field application. Entry in the field was retained successfully, validating the engineering applicability and promotion value of this technology. The primary conclusions are as follows.

The CRLD support accommodated the large roof deformation with high resistance during the retained entry service life. The roof on the DPB side was the focal supporting area, and the CRLD anchor there had an outstanding effect within 60 m behind the working face. DPB needed to be designed at a certain height and angle to make the gob roof collapse quickly and effectively. DPB application generated a directional crack and separated the roof between the retained entry and the gob. When the working face moved away, the gob roof caved in along the design position. The BGSS was designed into the anti-impact self-advancing structure, the sliding-yield structure, and the auxiliary supporting structure, which were well adapted to the mining pressure. The BGSS application integrated the gangue into an effective entry rib. The monitoring result showed that the gangue rib tended to be stable when the mining face advanced to a distance of 96 m.

The stable expanded gangue became the natural supporting body and decreased the mining pressure on the gob side. The retained entry under this pressure-relief circumstance did not deform much; it then entered the entry stabilization zone, where temporary support could be withdrawn. The onsite stabilization distance was 148 m. The gob-side entry was successfully retained in the field and non-pillar mining was realized. The retained entry quality is the critical factor in non-pillar mining technology. Adjusting the measures according to different mining geological conditions can improve the retained entry quality and make this technology universally applicable.

《Acknowledgements》

Acknowledgements

This work was supported by the National Key Research and Development Program of China (2016YFC0600900) and the Program of China Scholarship Council (201806430070).

《Compliance with ethics guidelines》

Compliance with ethics guidelines

Xingyu Zhang, Manchao He, Jun Yang, Eryu Wang, Jiabin Zhang, and Yue Sun declare that they have no conflict of interest or financial conflicts to disclose.