Research on the Movement of Overlying Strata in Shallow Coal Seams with High Mining Heights and Ultralong Working Faces

Abstract

To study the roof movement and ground pressure evolution characteristics of an ultralong working face in a shallow coal seam with a high mining height, the Shangwan Coal Mine in the Shendong mining area was used as the research background, and the physical and mechanical parameters of the surrounding rock were determined through rock mechanics experiments. A physical simulation model was built considering the 7 m mining height of the 12301 fully mechanized working face of the Shangwan Coal Mine to simulate and study the evolutions of the movement, fracture and collapse of the coal seam, direct roof, and basic roof and overlying strata during the mining process. The mechanical characteristics of the support, mechanism of roof collapse, and changes in the working resistance of the support were analysed and simulated. The research results indicate that when mining at a height of 7 m, the direct roof and basic roof strata collapse in layers; the basic roof strata collapse backwards, the rock block arrangement is more irregular, and the range of the basic roof that can form structural rock layers extends higher. After the basic roof rock fractures, it cannot form a masonry beam structure and can only form a cantilever beam structure. The periodic fracture of the cantilever beam causes periodic pressure on the working face. These research results are of great significance for planning the further mining of shallow coal seams with high mining heights and ultralong working faces in the Shendong mining area, as well as for improving the control of overlying strata.

1. Introduction

There are many shallow coal seams in northwest China, including Gansu, Ningxia, and Inner Mongolia. The burial depth of shallow coal seams is usually approximately 150 m, and such coal seams have typical characteristics, such as ultralong working faces, thin bedrock, and overlying thick loose sand layers [1,2,3]. The study of the mining pressure evolution at working faces with high mining heights is crucial for the efficient and safe mining of shallow and thick coal seams [4,5]. Based on the existing mining practice of fully mechanized shallow coal seam working faces with high mining heights in China, compared to the fully mechanized working faces of thin and moderately thick coal seams, the scale of goafs significantly increases [6]. After mining, the overlying roof rock layer fractures and fills a goaf, causing significant changes in the period, strength, and range of overlying rock movement and mining pressure in fully mechanized working faces with high mining heights [7]. Mining pressure problems caused by the high mining height and ultralong working face of shallow coal seams, such as roof collapse and roadway damage, seriously threaten the safety of personnel and equipment [8,9]. With increasing mining intensity and depth, the stress concentration and degree of damage intensify. Therefore, it is necessary to study the movement of overlying strata in shallow coal seams and the evolution of mining pressure distribution in high mining heights and ultralong working faces to ultimately achieve safe and efficient mining.

Many scholars have conducted extensive research on the development and evolution of mining pressure in shallow coal resources with different burial depths and working face lengths using theoretical analysis, physical simulation, and numerical simulation methods [10]. Li Yuanhui [11] used numerical and physical simulations to study the stress and failure characteristics of gently inclined thin ore bodies. As the dip angle of an ore body increases, the roof of the corresponding deeply buried mining area changes from experiencing stress release to experiencing stress concentration at different mining stages. When the dip angle of the ore body is 30°, shear failure is most likely to occur. Liu Chang [12] used a dual-mode parallel electrical method to dynamically monitor the overlying rock of a shallowly buried coal seam with a high mining face. Due to the influence of mining, cracks appeared in the near-surface loose layer above the working face. Zhang Libo [13] studied the spatial distribution characteristics and corresponding support design of the development and evolution of mining support pressure in the 8.2 m superhigh fully mechanized mining face with the 108 working face of the 12⁻² of Jinjitan Mine as the background, revealing the distribution characteristics of the dynamic stress field and internal and external stress fields. Li Ang [14] explored the safety and feasibility of deep coal seam floor mining in the North China Coalfield. Taking the coal seam mining in the Pingdingshan Coal Mine as the engineering background, FLAC^3D was used to simulate and analyse the influence of mining length, mining depth, and mining height on the depth of floor fractures. Ling Chunwei [15] noted that the hydraulic support in the mining faces of shallow coal seams with large mining outputs and strong rock strata around the stopes often resulted in damage. On the basis of basic experiments and physical similarity simulations, he studied the overlying rock fractures of shallow coal seams in the western mining area of China under the action of hydraulic coupling. Du Wengang [16] embedded distributed fibre optic and fibre optic grating sensors to study the deformation and internal stress evolution characteristics of overlying strata after shallow and thick coal seam mining and used them to monitor the deformation of overlying strata during coal mining. Zhang Cun [17] used drones, visible light cameras, and infrared cameras to determine the distribution characteristics and dynamic evolution of surface cracks. Based on the theory of rock movement and combined with monitoring results, a three-dimensional model of overlying rock fractures after mining in the 12,401 longwall working face was established. Liang Shun [18] studied the surface deformation during shallow coal seam mining and observed that as the working face depth increased, the surface deformation increased, reaching its maximum value near the coal seam. Zhang Jie [19] used theoretical analysis, numerical simulation, and similar simulation methods to study the evolution characteristics and formation mechanism of interlayer rock fractures during coal seam mining. Lan Tianwei [20] studied the mining pressure characteristics of shallow coal seams in extremely dense goafs and overlying coal–rock working faces and proposed distinguishing areas of the working face with similar mining stresses according to the advancing length of the working face, namely, the high-pressure zone, transition zone, and low-pressure zone. Niu Huiyong [21] reported that the gradual reduction in shallow coal seam resources, increase in coal seam mining depth, and increase in underground temperature have led to an increasing risk of residual coal fires each year.

The methods used in the study of thick and hard roof under the influence of mining mainly involve on-site measurement of mining pressure parameter data and analysis of mining pressure laws, similar simulation test research and inversion of overlying rock fracture structure morphology, numerical simulation calculation research and surrounding rock stress analysis, as well as construction mechanics models and theoretical analysis [22,23]. When using theoretical analysis for research, it mainly includes elastic–plastic mechanics, analysing thick and hard rock layers as rock slabs overlapped with the coal rock mass below, or the use of material mechanics to simplify the overlying rock as a rock beam model for research. Alternatively, based on structural mechanics, the stress analysis of rock blocks after the fracture of thick and hard rock layers can be conducted to study the criteria for structural instability after the fracture of thick and hard rock layers. By using similar simulation research, it is possible to analyse the morphology and evolution law of overlying rock fracture structures under the influence of mining from a macro perspective, and to more intuitively study the fracture characteristics of thick and hard rock layers under the influence of mining [24,25]. Numerical simulation calculation research is a research method that has emerged with the help of computer software. Based on failure criteria and models, it can efficiently and conveniently calculate and simulate the fracture mechanism of thick and hard rock layers under the influence of mining. The mining pressure data and parameters obtained from on-site engineering tests are the results of the fracture behaviour of thick and hard rock layers under the influence of mining. However, for the fracture characteristics of overlying rock structures, comprehensive analysis needs to be carried out in combination with the above theoretical research methods to verify the fracture situation of thick and hard roof slabs under the influence of mining.

The above research investigated the overlying rock failure in coal seams, the characteristics of crack development, the manifestation of shallow coal seam pressure, and effects of rock control measures. There is currently no systematic theory to guide the production of coal with high mining heights and ultralong working faces under shallow burial conditions. The movement of the overlying rock layers in the shallow coal seams of the Shangwan Coal Mine at ultralong and high mining heights was investigated. This article uses similar physical simulations to study the movement of rock layers due to shallow coal seam mining and establishes a theory of surrounding rock control for long working faces in shallow coal seams. The results of this modelling work can provide a theoretical reference for the subsidence of shallow coal seam strata during mining and has important theoretical and practical value for coal mine safety production, surface building protection, and rock movement research.

2. Geological Conditions of Coal Seams in the Shangwan Coal Mine

The Shangwan Coal Mine is located in the territory of the E’er’yijinhuoluo Banner, Inner Mongolia. The minable coal seams of the Shangwan Coal Mine include the #12 upper coal, #22 coal, and #31 coal seams from top to bottom. A total of 10 boreholes were surveyed in the 1⁻² coal seams of the Shangwan Mine, including 114, b214, sb10, sb11, sb12, sb15, b212, b213, Sh5, and W28. The rock parameters of these 10 geological boreholes are shown in

Table 1. Characteristics of the 1⁻² coal seams in the Shangwan coal mine.

Borehole	Burial Depth/m	Alluvium/m	Bedrock/m	Immediate Roof/m	Thickness/m	Main Roof/m	Thickness/m	Coal Thickness/m
114	118.83	8.2	110.63	Sandy mudstone	7.11	Coarse sandstone	9.25	4.68
B214	84.32	11.28	73.04	Sandy clay	0.55	Fine and siltstone	4.56	7.1
Sb10	125.62	11.7	113.92	Sandy mudstone	0.6	Coarse sandstone	9.89	6.86
Sb11	110.25	13.88	96.37	Sandy mudstone	1.22	/	0	7.5
Sb12	120.8	20.93	99.87	Mudstone, sandy mudstone	1.71	Fine sandstone	5.83	5.88
Sb15	97.41	10.33	87.08	Sandy clay	5.62	Coarse sandstone	13.89	3.8
b212	88.77	0.8	87.97	Fine sandstone	5.49	Middle sandstone	15.11	9.25
b213	98.82	23.85	74.97	Fine sandstone	1.25	Coarse sandstone	12.41	8.64
Sh5	165.04	16.73	148.31	Sandy mudstone	1.03	Coarse sandstone	9.55	6.63
W25	144.8	15.98	128.82	Sand clay	0.6	Middle sandstone	11.95	9.62
Average	115.47	13.37	102.1		2.52		9.24	7.00

. The 1⁻² coal seams in the Shangwan mining area have burial depths ranging from 84.32 to 165.04 m, with an average of 115.47 m. The thickness of the alluvial sand varies from 0.8 to 23.85 m, with an average thickness of 13.37 m. The thickness of the bedrock is 73.04~148.31 m, with an average thickness of 102.1 m. The direct roof lithology is mostly sandy clay, sandy mudstone, or siltstone, with a thickness range of 0.55–7.11 m and an average thickness of 2.52 m. The basic roof lithology is mostly coarse sandstone and fine sandstone, with a thickness of 0–15.11 m and an average of 9.24 m. The thickness of the 1⁻² coal seams in the Shangwan mining area ranges from 3.8 to 9.25 m, with an average thickness of 7 m.

From bottom to top, sandy mudstone, fine-grained sandstone, siltstone, fine-grained sandstone, siltstone, medium-grained sandstone, and siltstone is present. The coal seam spacing is small, and the overlying rock layer has a greater strength. A stratigraphic column of the coal seams is shown in Figure 1.

3. Similarity Simulation Experiment

3.1. Experimental Design

According to the prototype conditions, a large planar strain testing device with a length × height × thickness of 3000 mm × 3000 mm × 200 mm was selected for laboratory testing [26,27]. The surroundings and bottom plate of the test bench are constrained by 20# channel steel and a 25 mm thick organic glass plate, with the top of the model being the free end. The studied coal seam is buried at a depth of 118.76 m and has a thickness of 7 m. Based on the prototype conditions and experimental setup, the geometric similarity ratio is determined to be C_l = 1/50, the bulk density similarity ratio is C_γ = 17/25 = 0.68, the stress similarity ratio is C_α = C_l × C_γ = 0.68/50 = 0.0136, and the time similarity ratio is $C_t=\sqrt{C_l}=0.1414$.

The model has a length of 3000 mm, with 200 mm wide coal pillars left on both sides of the model. The model can simulate a working face advancement length of 2600 mm, which can reflect a real working face advancement length of 130 m. The model is 200 mm thick and can simulate a working face length of 10 m. In the model, the thickness of the coal seam floor is 208.6 mm, the thickness of the coal seam is 140 mm, the thicknesses of the coal seam roof stratum and aeolian sand are 2375.2 mm, and the total height of the model is 2723.8 mm. Based on the coal seam thickness and geological conditions in the Shendong mining area, a planar similarity simulation test bench was used to simulate the development of mining pressure and the evolution of the overlying rock movement during the mining process of a full-thickness working face with a coal seam thickness of 7 m.

Based on the similarity theory, experiments are conducted using similar simulation materials to simulate the hinge structure morphology and migration evolution process of overlying fractured rock blocks during underground coal and rock excavation construction. Essentially, it is based on the principle of similarity, calculating and reducing the physical quantities of underground mine rock layers according to similarity ratios, preparing similar materials, and establishing models. Then, in the model of similar simulation research, simulating underground coal and rock excavation construction according to the pre-designed plan can provide a visual analysis of the evolution process and failure-instability-induced disaster mechanism of the hinged structure of overlying rock blocks under the influence of strong mining from a macro perspective of the mining site. The micro and internal damage situation cannot be displayed, and the model is a simplified model, which has certain differences from the on-site production situation. The similarity phenomenon refers to the proportional relationship between geometric-related physical quantities and their corresponding instantaneous changes being a certain value, and the existence of the same physical laws. We refer to the ratio of the same dimensional physical quantity at the corresponding point in a similar process as the similarity coefficient, also known as the similarity constant. The similarity constant exists in a certain combination relationship in similar phenomena. Substituting the same physical quantities in the similarity index into the same system can generate dimensionless combinations of various physical quantities. We call this combination the similarity criterion. Calculate the geometric similarity ratio based on the dimensions of the prototype’s working face and experimental setup.

3.2. Model Materials and Proportions

Considering the discontinuity of rock materials and the measured mechanical strengths of rocks, a crack influence coefficient of 0.8 was considered [28,29]. Based on the mechanical parameter data of on-site rock testing on the working face and referencing other relevant literature on geological rock testing on the working face, uniaxial compression experiments were conducted on standard specimens made in the laboratory, and the optimal ratio of similar materials through repeated adjustments were obtained. Then, based on the similarity theory, the physical and mechanical parameters of each rock layer in the model were obtained, the material ratio was selected, and the material dosage was calculated. For details on the model parameters, please refer to

Table 2. Main parameters of the roof and floor of the simulated mining coal seam and proportions of materials adopted in the model.

Layer	Lithology	Layer Thickness /m	Density /(kg/m3)	Compressive Strength /MPa	Elastic Modulus /GPa	Similarity Ratio	Sand /kg	Cement /kg	Calcium Carbonate /kg	Gypsum /kg	Water /kg	Borax /g
1	Windblown sand	11.70	1700	/	/	/	238.7	/	/	/	/	/
2	Siltstone	8.60	2460	40.6	35	655	152.9	/	12.3	12.3	33.6	336
3	Sandy mudstone	5.10	2240	22.8	23	346	74.6	9.3	/	14	9.3	186
4	Coarse sandstone	10.28	2430	36.6	35	855	185.7	/	11.3	11.3	20.4	204
5	Fine sandstone	7.03	2500	44.6	32	955	131.9	7.2	/	7.2	14.3	143
6	Sandy mudstone	9.8	2240	22.8	23	337	143.3	/	13.4	31.4	25.6	512
7	Coarse sandstone	11.24	2430	36.6	35	855	203.1	/	12.4	12.4	22.3	223
8	Sandy mudstone	27.52	2240	22.8	23	337	402.4	/	37.7	88.0	71.9	1437
9	Coarse sandstone	9.89	2430	36.6	35	855	178.7	/	10.9	10.9	19.6	196
10	Sandy mudstone	6.17	2240	22.8	23	337	90.2	/	8.5	19.7	16.1	322
11	Fine sandstone	7.27	2500	44.6	32	955	136.4	7.4	/	7.4	14.8	148
12	Sandy mudstone	4.16	2240	22.8	23	337	60.8	/	5.7	13.3	10.9	217
13	Coal seam	5.5	1480	10.5	15	373	53.1	11.6	/	5	6.6	133
14	Siltstone	1.48	2460	40.6	35	655	26.3	/	2.1	2.1	4.2	42
15	Sandy mudstone	3.25	2240	22.8	23	337	47.5	/	4.5	10.4	8.5	170
16	Fine sandstone	5.70	2500	44.6	32	955	107	5.8	/	5.8	11.6	116

.

3.3. Model Testing

The Nikon DTM-531E total station was used to measure the coordinates of the measurement points arranged in the direct roof, basic roof, and overlying rock layers of the coal seam to obtain the displacement of the model. A CM-2B static strain gauge was used to collect the strain of the stress box pre-embedded inside the model. The stress box was made of a BE120-3AA resistance strain gauge, a custom-made copper ring, and a copper sheet. The calibration of the stress box was carried out before the test. The strain of the stress box measured by the strain gauge in the experiment was calculated via the calibration data. The stress box is shown in Figure 2, and the stress testing system is shown in Figure 3. A digital camera was used to record typical phenomena during the mining process.

(1)

Layout and testing methods of the rock movement measurement points

The displacement measurement points inside the rock layer can be conveniently used to monitor the movement of the rock layer, as shown in Figure 4. Thirteen rows of measurement points are arranged horizontally on the overlying strata of coal seam 1–2, with a spacing of 20 cm. The survey points were arranged 5 cm above the coal seam in the vertical direction. The first four rows had a spacing of 10 cm, and the last three rows had a spacing of 15 cm. A total of 96 displacement measurement points were arranged.

(2)

Layout and testing methods of pressure measurement points

In the model experiment, to determine the distribution of the support pressure in front of the working face during the mining process, stress sensors were installed in conjunction with static and dynamic strain gauges, as well as a computer automatic data acquisition system for model stress measurement. Two rows of BE120-03AA (11) resistance strain gauges were buried with the stress sensors, 5 cm, 15 cm, and 25 cm from the simulated mining coal seam roof. The stress of the direct roof and basic roof during the mining process was measured. The spacing between points was 20 cm. The layout of the stress observation sensors for simulating the mining of coal seams 1–2 is shown in Figure 5.

4. Analysis of the Similar Experiment Results

4.1. Experimental Process Description

After the model was made and dried for 5 days, the glass panel was opened for displacement measurement point arrangement, as shown in Figure 6. The base point for the total station measurement was selected, the initial readings of each measurement point were taken, and the pressure box and strain gauge were connected, as shown in Figure 7.

A cut was made at a distance of 10 m from the edge of one side of the model, with a width of 8 m and a height of 5.5 m. A four-column hydraulic support was placed inside the cut. The underground work was carried out according to the 38th working system, with two-and-a-half shifts for coal cutting and half-shifts for maintenance. The coal cutting time for each knife cut in the actual scenario is approximately 1.3 h, and the advance distance is 0.8 m. Therefore, each shift cuts 6 coal seams and 15 coal seams per day. According to the time and geometric similarity ratio, the coal cutting time for each knife cut in the experiment was 11 min, and the advance distance was 0.8 m, with a total progress of 12 m per day. The relationship between the distance and cutting process is shown in

Table 3. Distances and cutting process.

Distances/m	Cutting Process	Distances/m	Cutting Process	Distances/m	Cutting Process
15.2	19	40	50	84.8	106
17.6	22	50.4	63	95.2	119
19.7	24	55.2	69	100	125
20	25	64.8	81	110.4	138
29.6	37	69.2	86.5	115.2	144
35.2	44	80	100

.

Figure 8 shows the evolution process of the first collapse of the direct roof during the advancement of the working face. When the working face was advanced to 15.2 m (9th cut), the first and second layers separated. When the working face was advanced to 17.6 m (12th cut), directly above the first layer of collapse, there was a collapse thickness of 1 m. Separation occurs between the second and third layers. When the working face was advanced to 19.7 m (the 14th cut), the second layer collapsed, and the third layer collapsed in blocks. The total thickness of the collapse was 3 m, and the remaining height of the goaf space was 4.5 m. When the working face was advanced to 20 m (15th cut), all the collapsed roof moved. The goaf was directly filled without causing collapse, with a maximum free space height of 3.6 m. The direct roof collapsed completely, and the basic overhanging length of the roof was 10 m. The direct collapse angle (on the side near the coal pillar) was 52°.

Figure 9 shows the evolution process of the initial collapse of the basic roof during the advancement of the working face. When the working face was advanced to 29.6 m (the 27th cut), the first and second layers of the basic roof began to collapse. When the working face was advanced to 35.2 m (the 34th cut), the second layer of the structure formed by collapse experienced sliding instability. When the working face was advanced to 20 m, the third and fourth layers of the basic roof collapsed. At 50.4 m (53rd cut), the fifth layer of the basic roof collapsed, causing the overlying rock layer to collapse. The working face was under initial pressure, and when the working face was advanced to 50.4 m, a crack with a length of 20 m appeared at the top of the third layer at 8.5 m in front of the working face.

When the working face was advanced to 55.2 m in Figure 10, the cracks in front of the working face extended upwards and downwards, penetrating directly the rock above the support. When the working face was advanced to 64.8 m, the cracks in front of the working face continued to extend. New fracturing cracks appeared directly above the support. When the working face was advanced to 69.6 m, with various cracks developing and the basic roof breaking, the overlying rock layer collapsed and caused the first periodic pressure on the working face. After being left for 12 h, cracks appeared at a height of 12.5 m above the fifth layer in front of the working face at a height of 15 m. When the working face was advanced to 75.2 m, the crack in the working face extended downwards to the third layer of the basic roof, with a length of 25 m. When the working face was advanced to 80 m, the cracks continued to expand downwards to the direct roof, reaching 50 m upwards. The subsidence of the top plate increased, and the second cycle of pressure developed.

When the working face was advanced to 84.8 m, after the removal of the support, the roof rock layer was cut off along the back of the support, and the second cycle of pressure was completed. When the working face was advanced to 95.2 m, a crack with a length of 17.25 m above the coal wall formed, indicating the start of the third cycle of pressure. When the working face was advanced to 100 m, the third basic roof fracture caused periodic pressure with a step distance of 20 m. When the working face was advanced to 110.4 m, a crack approximately 20 m long appeared in the fourth and fifth layers above the coal wall. When the working face was advanced to 115.2 m, cracks expanded above the coal wall, the roof sank, and the fourth cycle of pressure was applied to the working face in Figure 11.

4.2. Roof Collapse Evolution

(1)

Direct roof: First collapse step distance

When the working face was advanced to 15.2 m, there was separation between the first and second layers. When the working face was advanced to 17.6 m, the rock directly above the first layer collapsed, with a collapse thickness of 1 m. Separation occurred between the second and third layers. When the working face was advanced to 20 m, the entire roof collapsed. The goaf was directly filled without collapsing, with a maximum free space height of 3 m. The direct roof collapsed completely, and the basic overhanging length of the roof was 10 m. The direct collapse angle (on the side near the coal pillar) was 52°.

(2)

Basic roof: First collapse step distance

When the working face was advanced to 29.6 m, the first and second layers of the basic roof began to collapse. After the second layer fractured, the two pieces interlocked to form a three-hinged arch structure. At 40 m, the third and fourth layers of the basic roof collapsed. At 50.4 m, the fifth layer of the basic roof collapsed, and the overlying rock layer collapsed accordingly. The working face was under initial pressure, with a step distance of 50.4 m in

Table 4. Statistics of the immediate roof and basic roof collapse intervals.

Collapse Interval	Pressure	Distance of the Working Face Advance under Pressure		Loading Step
		Model/mm	Prototype/m	Model/mm	Prototype/m
1	First direct roof collapse	400	20	400	20
2	Basic roof initial pressure	1008	50.4	1008	50.4
3	First cycle of pressure	1392	69.6	384	19.2
4	Second cycle of pressure	1600	80	208	10.4
5	Third cycle of pressure	2000	100	400	20
6	Fourth cycle of pressure	2304	115.2	304	15.2
7	Mean cycle pressure step distance			324	16.2

.

(3)

Periodic loading step distance

After experiencing the initial pressure, as the working face continued to advance, the structure formed by the basic roof strata always underwent a change from stable to unstable and then back to stable, which was a cyclic process called periodic pressure [30,31]. During the excavation process of the model, the working face underwent four cycles of pressure, with pressure steps of 19.2 m, 10.4 m, 20 m, and 15.2 m.

4.3. Movement Pattern of the Overlying Rock Layers

A displacement measurement point was arranged in the middle of the direct roof, with the number of measurement points ranging from #1 to #12. A measurement point slowly sank after a hole was cut from the working face. As the working face was advanced, layer separation occurred in the roof, and the displacement of the measurement point increased sharply. When the working face was advanced to 20 m, displacement measurement point #1 collapsed directly. When the working face was advanced to 35.2 m, displacement measurement point #2 collapsed directly. When the working face was advanced to 50.4 m, displacement measurement point 3 directly collapsed. The corresponding displacement curves are shown in Figure 12.

Through the analysis of the relationship curves between the settlement of displacement measurement points #1, #2, and #3, and the advancing distance of the working face, it can be concluded that before direct roof collapse, the direct roof generally experienced delamination and settlement, usually approximately 1–2.5 m. The amount of sinking of the collapsed roof was between 3.5 and 4.5 m. As the working face continued to advance, the amount of sinking of the direct roof under the pressure of the overlying rock layer gradually grew to approximately 5 m and then stabilized.

4.4. Basic Roof Subsidence and Collapse

Two layers of displacement measurement points in the middle and upper parts of the basic roof were arranged to measure the settlement of the basic roof [32,33]. The measurement points are numbered from #13 to #36. The variation curves of the subsidence of measurement points #13~15 and #25~27 with respect to the advancing distance of the working face are shown in Figure 13. The basic roof collapsed in layers. When the working face was advanced to 35 m, the lower layer measurement points collapsed when the basic roof first collapsed, while the upper layer measurement points collapsed when the basic roof collapsed only approximately 50 m. The settlement of the upper and lower displacement measurement points before collapse was between 1 and 1.5 m.

4.5. Subsidence and Collapse of Rock Layers

There were five layers of displacement measurement points buried in the rock layer above the basic roof, with the first layer 100 mm from the measurement point above the basic roof and the other four layers spaced 150 mm apart. The measurement points at the third and sixth layers were blocked by channel steel, making it impossible to measure the data. The selected displacement measurement points were 1400 mm from the model boundary and are numbered #43, #55, #79, and #91. The relationship curve between the settlement of the measurement points and the advancing distance of the working face is shown in Figure 14. Before the working face was advanced to 65 m, the subsidence of the rock layer above the basic roof was very small. Later, under the pressure of the roof, tensile cracks appeared in the overlying rock layer in front of the working face, and the subsidence of the displacement measurement point gradually increased. When advancing to 80 m, the sinking speed of the displacement measurement point increased, and it entered the second cycle of pressure on the basic roof. The rock layer above the basic roof collapsed with the collapse of the basic roof and was finally compacted. The sinking amount of the displacement measurement point tended to stabilize [34].

4.6. Distribution Characteristics of Support Pressure

The maximum stress at the first-layer stress measurement point occurred an average distance of 12.5 m from the working face, with an average stress of 7.68, an average stress concentration coefficient of 2.64, and an average range of influence of the support pressure of 32.5 m. The maximum stress at the second-layer stress measurement point occurred 8.75 m from the working face on average, with an average stress of 7.62, an average stress concentration coefficient of 2.74, and an average range of influence of the support pressure of 32.5 m. The maximum stress at the third stress measurement point occurred an average distance of 5 m from the working face, with an average stress of 7.71, an average stress concentration coefficient of 2.9, and an average range of influence of the support pressure of 35 m. According to the stress measurement point data, as the layer of the roof rock at the stress measurement point increased, the maximum support pressure approached the working face, and the stress magnitude, stress concentration coefficient, and influence range did not substantially change. The maximum stress at the third-layer stress measurement point was 8.75 m from the working face on average, with an average stress of 7.67, an average stress concentration coefficient of 2.76, and an average range of influence of the support pressure of 33.3 m.

The overlying rock in the mining area experienced the influence of repeated mining on the previous adjacent working face and this working face before and after mining. After mining, the thick and hard rock layer in the goaf broke, and the adjacent fractured rock blocks formed a masonry rock beam structure under the interlocking effect. After the coal seam in the lower working face is mined out, an unloading space is formed in the overlying rock of the roof above the goaf. As the coal seam is gradually mined, when the unloading space range meets certain conditions, the overlying rock layer will collapse under the action of its own weight and overlying rock load [35,36]. The overlying rock collapse pattern is a transitive development trend from the bottom to the top of the coal seam roof, with obvious layering and grading characteristics. The activity of the overlying strata in the goaf is the result of the interaction between the stress in the mining area and the strength of the overlying strata. When the roof overlying strata collapse to the position where the ultimate bearing rock layer is located, their collapse height no longer develops. In the overlying strata in the goaf, the structural effect of “stress arch” is formed due to the self-organizing limitation of the overlying strata failure and movement. At the same time, the interaction between the fractured blocks after the collapse of the overlying rock in the mining area forms the “fractured arch”; the “fractured arch” in the overlying rock structure of the mining site is located within the range of the “stress arch” pressure relief zone. The stress arch mainly bears and transmits the load of the overlying rock layer; the hinge structure formed by the fracture of the overlying rock in the “fractured arch” mining area is the main load factor for the manifestation of strong mining pressure.

5. Discussion

Through the observation of mining pressure in the practice of mining working faces with high heights and ultralong lengths, as well as research on existing theories of mining pressure and rock layer control, it is shown that changes in the length of the working face will lead to changes in the roof pressure during the mining process, which indicates that the height of the roof rock layer movement and collapse in the mining area changes, and this change conforms to the pressure balance arch in the Proctor theory. The Proctor theory suggests that the roof strata above a certain span space (working face length) will experience arch collapse, and the height of the arch is related not only to the collapse but also to the strength of the rock. The specific relationship is shown in Formula (1). The height of the roof rock caving is

$$H_{M,L_{\mathrm{g}}}=\frac{L_{\mathrm{g}}}{2\mathrm{f}}$$

(1)

where $H_{M,L_{\mathrm{g}}}$ is the falling height, m; $L_{\mathrm{g}}$ is the working face length, m; and f is Proctor’s coefficient.

5.1. Theoretical Prediction of the Impact of Changes in Working Face Length on Support Strength

The collapse height of the roof rock layer under different working face lengths can be determined, and therefore, the mining pressure induced during the unloading period of the working face can be obtained as

$$q_2=\frac{\gamma •{\mathrm{L}}_{\mathrm{g}}}{2\mathrm{f}}$$

(2)

where q₂ is the induced mining pressure in the normal working face, kPa, and γ is the rock bulk density, kN/m³.

There are many coal seams with thicknesses greater than 6.3 m and a geological reserve of 1351.99 Mt in the Shendong mining area. The current problem faced by the Shendong mining area is what kind of coal mining method to adopt to achieve scientific, economic, safe, and efficient mining and to maximize the recovery rate of the coal resources. In general, there are three main mining methods for thick coal seams: layered mining, top coal caving mining, and full-height mining. The following work analyses and demonstrates the advantages and disadvantages of these three methods and their applicability to the Shendong mining area from different perspectives.

5.2. Analysis of the Applicability, Advantages, and Disadvantages of Layered Mining in Extrathick Coal Seams

(1)

Layered mining is technically feasible

For the mining of thick coal seams, layered mining was mainly adopted in the 1970s and 1980s. Many experts and scholars have conducted comprehensive research on the technical issues in layered mining of thick coal seams, making it a mature and feasible technology. When mining thick coal seams in layers, metal mesh must be laid during the mining process of the first layer to form a false roof and facilitate the mining of the lower layer. The hanging mesh method for long-wall layered mining of thick coal seams was studied for different roof types, and good results were achieved via trial. A study was conducted on the movement of overlying strata in the layered mining of ultrathick coal seams. Due to the increase in total mining height, the movement of rock layers in the fractured zone is different from that in the curved subsidence zone. The rock layers in the curved subsidence zone still follow the sinking trend of masonry beams, while the rock layers in the fractured zone move in a stepped manner. The horizontal distance of the steps is suitable for the pressure step distance in the mining area. An experiment using a cemented roof to mine thick coal seams layer-by-layer showed that after the underground solidification period of 3 months, the time for the cemented roof material to reach its maximum strength is 28 days. Thus, it is best to use cover-type supports at the mining face.

(2)

Layered mining has poor applicability in terms of safety, efficiency, production, and economy

For the 7 m thick coal seam in the Shendong mining area, if layered mining technology is adopted, the impact on the high production and efficiency of the mine and the economic efficiency of coal seam mining mainly includes the following aspects.

①

Large amount of excavation work and maintenance costs for layered mining tunnels

The 7 m thick coal seam adopts a layered mining method, which requires at least two layers of mining. Therefore, the amount of excavation work for the tunnel needs to be doubled. Moreover, the roof of the lower layer roadway is prone to fragmentation due to the compaction of the collapsed gangue in the upper layer, resulting in an increase in maintenance costs for the roadway.

②

The increasing number of moves in the layered mining face makes it difficult to achieve high productivity and efficiency.

For the 7 m thick coal seam, layered mining requires at least two layers of mining, so the layered mining face needs to be moved at least twice, and frequent movement of the face will lead to a waste of time and difficulty in achieving high production and efficiency.

③

Low recovery rate of layered mining

The practice of layered mining in thick coal seams has shown that for the Shendong 7 m thick coal seam, layered mining is generally divided into two layers, and a top coal layer of 0.5–1.0 m must be left between the layers to ensure the stability, safety, and control of the roof of the machine passage space during mining. Leaving the top coal will decrease the recovery rate of the coal seam.

④

Increased risk of spontaneous combustion of coal remaining in the goaf during layered mining

The layered mining process requires leaving a portion of the top coal to maintain the stability of the roof during lower layer mining. This portion of the coal is left in the goaf after the lower layer of the working face is mined, which poses a risk of spontaneous combustion. Therefore, the cost of fire prevention during the mine production process will definitely increase.

The above analysis shows that the use of layered mining in a 7 m thick coal seam is technically feasible, but it will increase coal mining costs, increase risks, and make it difficult to achieve high production and efficiency.

5.3. Analysis of the Applicability, Advantages, and Disadvantages of Fully Mechanized Caving Mining in Extrathick Coal Seams

The coal seams in the Shendong mining area are characterized by shallow burial, thin bedrock, moderate hardness, and high toughness. These conditions cause fully mechanized top coal caving mining to be ineffective; the recovery rate is low, and it is difficult to achieve high production and efficiency.

①

Shallowly buried coal seam conditions (50–200 m) result in a poor roof coal caving ability

Under the conditions of shallowly buried coal seams, the support pressure is low, and the top coal cannot be fully broken under the support pressure. Therefore, if top coal caving is used to extract the extra-thick coal seam in the Shendong mining area, the block size of the falling top coal during mining is large, and the caving performance is poor.

②

The conditions of the moderately hard and high-toughness coal seams in the Shendong mining area make it difficult to ensure the top coal release rate

The coal seams in the Shendong mining area are Jurassic coal seams with a uniaxial compressive strength of 20–30 MPa and high toughness. The crushing and caving characteristics of the top coal are closely related to its uniaxial strength. The difficulty in ensuring the top coal release rate leads to a lower recovery rate of the entire coal seam.

Considering the technical conditions and high production and efficiency of the Shendong extra-thick coal seam, fully mechanized top coal caving mining cannot adapt to the mining of the 7 m extra-thick coal seam.

5.4. Applicability Analysis of High Mining Height in Ultrathick Coal Seams

In recent years, the mining height in the Shendong mining area has increased from working face heights of 4.5 m to 5.5 m and 6.3 m, and the mining technology used to achieve high mining heights in this mining area is state-of-the-art within China and even worldwide.

(1)

Large-mining-height technology can adapt to the mining of thick coal seams in the Shendong mining area

There have been many successful production practices in the Shendong mining area using high-mining-height technology, with rich technical experience. Moreover, the heights of the large working faces in the Shangwan and Bulianta mines reached 6.3 m. With the study of practical rock control and coal mining technology at a height of 6.3 m and scientific research and key technology development for a height of 7 m, it is possible to carry out high mining of the 7 m thick coal seam in the Shendong mining area while ensuring safety, high production, and efficiency.

(2)

The Shendong mining area has an extremely thick coal seam, and high-mining-height technology is economically optimal

The main factors affecting the economic viability of the Shendong mining area’s ultrathick coal seam, if high-mining-height technology is used, are as follows:

①

The amount of excavation work for high-mining-height mining tunnels is small

Compared to layered mining and fully mechanized top coal caving mining, mining extra-thick coal seams with working face heights greater than 7 m requires less excavation and support at high mining heights. The tunnel system is simple, and excavation replacement is easier to coordinate, which is conducive to the high yield and efficient mining of extra-thick coal seams.

②

Low frequency of movement during high-mining-height mining

For a coal seam with a thickness of 7 m, an area of coal seam is moved only once at full height, with the lowest cost.

③

High coal seam recovery rate during high-mining-height mining

The Shendong mining area uses a 7 m thick coal seam and hydraulic supports with appropriate support heights. High-mining-height mining technology can extract most of the resources from thick coal seams, and the recovery rate of the working face can reach over 95%.

Based on the above analysis, high-mining-height mining can achieve a high yield and efficient mining of 7 m thick coal seams both technically and economically.

6. Conclusions

(1)

This study was based on the parameter information of rock pressure manifestation in the surrounding rock of the fully mechanized caving face with thick and hard roof under mining in Shangwan Coal Mine. The manifestation law of mining pressure induced by the fracture of thick and hard roof under mining was clarified. When mining at a height of 7 m, the initial collapse step distance of the direct roof is 20.8 m, and the initial collapse step distance of the basic roof is 44.8 m. During the mining simulation process, a total of six cycles of pressure are experienced, with an average step distance of 12.9 m.

(2)

Based on the manifestation law of mining pressure induced by roof fracture under mining, this paper reveals the time course required for the development, penetration, and instability of thick and hard rock strata cracks in the working face under strong mining, and elaborates on the temporal effect and regional characteristics of the development and evolution of rock strata cracks under mining. Before direct roof collapse, the amount of direct roof subsidence is generally approximately 0.5–2.5 m, while after collapse, the amount of direct roof subsidence is 3.5–5 m. As the working face continues to advance, the amount of direct roof subsidence under the pressure of the overlying rock layer gradually increases to approximately 5–7 m and then stabilizes.

(3)

Based on the fracture structure and evolution law of a thick and hard roof under strong mining, the energy storage mechanism of a thick and hard roof under strong mining and the main control factors causing disasters are elucidated. When mining at a height of 7 m, the advanced support pressure in front of the coal wall has a certain distribution. From the analysis of the data measured by the stress sensor in the experiment, the average distance between the maximum stresses of the first-layer stress measurement point and the working face is 9.25 m, the average stress is 7.45 MPa, the average stress concentration coefficient is 2.56, and the average range of influence of the support pressure is 30.75 m.