Overburden Breaking Law and Safe Mining Technology in Thin Bedrock Stope with Thick Alluvium

Abstract

For the thick alluvium and thin bedrock coal seam mining in East China, it is easy to produce water by pressing the frame. Taking the engineering geological conditions of the thin bedrock coal seam 1611 (3) working face in the Zhangji Coal Mine of Huainan Mining Group as the research background, the breaking law of the thin bedrock coal seam working face with thick alluvium was studied by using a similar material physical simulation experiment and theoretical analysis. It is revealed that under the load transfer of the loose aquifer, the roof fracture of the stope with thick alluvium and thin bedrock is characterized by large and small periodic ground pressure, and the combination failure of the high key stratum and the low key stratum with large cycles is prone to support crushing accidents. Based on this, the safety mining technical measures of reducing the fracture distance of the high key stratum by forced roof caving are proposed, and the reasonable blasting parameters are calculated. After field application, no support crushing accident on the working face was recorded, and the mining was safe, which provides a technical reference for the mining of the working face under similar conditions.

1. Introduction

There is a loose aquifer composed of unconsolidated sandy soil and sandy gravel rock as the skeleton at the bottom of the Quaternary thick overburden of many coal mines in East China, North China and other mining areas in China [1]. It is the fourth aquifer (commonly known as “four aquifers”) in the overburden which is directly located on the top interface of the coal measure bedrock. Experience proves that the loose aquifer is no safety hazard to coal seam mining when the waterproof coal pillar of a certain thickness is set normally [2,3]. However, in recent years, when mining the coal seam near the waterproof coal pillar, and the working face is near the loose aquifer, there have been a lot of disaster accidents caused by stope pressure of fully mechanized working face. Therefore, it has attracted the attention of industry decision makers and many scholars. Experts and scholars at home and abroad have conducted in-depth research on the safe mining of thin bedrock coal seam. According to the literature [4,5,6], the maximum caving height and effective water diversion height of overburden under different mining and caving conditions are obtained, and the fracture development height of thin bedrock is detected on site. According to the literature [7,8,9], the shallow buried thin bedrock coal seam was studied. The shallow buried coal seam was defined as the coal seam with a burial depth no more than 150 m and a base-mining ratio of less than 1, and the ore pressure characteristics of the shallow coal seam were obtained. The roof acted on the working face in the form of integral cutting, and the structural instability conditions of the key stratum were analyzed. Based on our updated literature survey, researchers, e.g., [10,11], studied the law of ore pressure in the fully mechanized caving face of the Sima Mine with thin bedrock in the Lu-an Mining area, and obtained the law of mining-induced fracture development and overburden fracture mechanism. The literature cited in [12,13,14,15], based on the research background of coal mines in the Huainan and Huaibei mining areas, has obtained the influence of mining factors such as the structural characteristics of overburden in thin bedrock working face and the base mining ratio on the migration law of overburden, and studied the mechanism of water inrush and support crushing accidents. To sum up, previous studies mainly focused on shallow–deep thin bedrock coal seam, alluvial thickness of 100–300 m, and mainly focused on the size of waterproof safe coal pillar, the development height of water-conducting fracture zone and the disaster mechanism of shallow buried thin bedrock ballast, etc. However, there are few or insufficient studies on the ore pressure in the deep buried thin bedrock stope with an alluvium thickness of more than 400 m in the Huainan and Huaibei mining areas. Taking the 1611 (3) working face of the Zhangji Coal Mine of Huainan Mining Group as the engineering background to increase the upper limit of mining, the thickness of the overlying alluvium is about 440 m. Our paper studies the characteristics of overburden fracture and the coal seam safety mining technology when mining the deep buried thin bedrock coal seam under this occurrence condition, providing theoretical guidance and reference for the safe mining of the working face under similar conditions.

2. Engineering Geology Overview

The 1611 (3) working face of the Zhangji Coal Mine of Huainan Mining Group is the first working face of 13-1 coal in the Xisan Mining area. The elevation of the working face is (−400)–(−460) m, which belongs to the working face category with a higher mining upper limit, as shown in Figure 1. The incline length of the inner cutting face is 117 m and the strike length is 170 m. The bar chart of the drill hole near the inner cut is shown in Table 1. According to the calculation method in the literature [6], it can be calculated that the fine sandstone at the depth of 424.7 m is the low key stratum, and the medium-fine sandstone at the depth of 406.8 m is the high key stratum. The average thickness of the coal seam is 6.1 m, the structure is complex, the average dip angle of the coal seam is 11.5°, and the initial mining height of working face is 5 m.

3. Experimental Study of Overburden Migration Law in Thin Bedrock Stope

3.1. Subsection

In order to find out the migration law of overburden during the mining of deep-buried thin bedrock coal seam, our paper takes the 1611 (3) lifting face as the engineering geological background to carry out a physical simulation test of similar materials. The length, width and height of the model frame used in the test are 3000 mm × 300 mm × 1200 mm, respectively, and the geometric similarity ratio of the test was 1:100. The mechanical parameters of each rock layer are given in Table 2. According to the method of reference [12], the physical simulation experimental model of similar materials was established, as shown in Figure 2a. The water–sand belt simulates the aquifer, and the water pressure equalization system is connected on both sides of the water–sand belt (Figure 2b) to simulate the flow recharge characteristics of the aquifer. Two stress measuring lines are arranged in the intermediate strata between the coal seam and the loose aquifer to test the change law of stress during mining.

3.2. Migration Characteristics of Overburden in Stope

The mining process of the working face is shown in Figure 3.

According to Figure 3, when the working face advances to 50 m, the stope’s initial pressure can be seen in Figure 3a. With the increase of stope distance, the stope is subject to the first cycle to pressure. When the working face advances to 70 m, the second cycle of pressure takes place on the stope, and the pressure step distance is 12 m. When the working face advances to 85 m distance, the upper key stratum breaks and collapses, while the lower key stratum also breaks accordingly. It can be seen that the upper key stratum and the lower key stratum are broken together during the pressing this time (Figure 3c), which is the first large cycle pressing in the stope. In the subsequent mining process, the working face pressure also has the characteristics of cyclic pressure. When the working face advances to 130 m, the stope roof appears to have the second big cycle caving (Figure 3e), and the caving step is 45 m.

It can be seen from Figure 4 that the overburden height of fracture subsidence increases during the model mining process. When the working face advances to 70 m, the overburden fracture height value is 20 m above the coal seam roof (Figure 3b). When the working face advances to 85 m, the height of overburden breaking subsidence abruptly increases to 40 m above the roof of coal seam, and the range of transverse breaking subsidence also increases substantially (Figure 3c), indicating that the key stratum above the stope is broken and subsiding, and the roof of the working face has the first major period of pressure. When the working face advances to 120 m, the movement height of overburden fault does not change, but the strata above the coal seam roof within 15 m are broken and subsiding (Figure 3d), indicating that the key stratum in the lower part repeats small periodic breaking at this time. When the working face advances to 130 m, the overburden height of the fracture subsidence is 45 m, but the overburden fracture displacement height near the working face suddenly increases (Figure 3e), indicating that the key stratum of the upper part of the working face has another fracture subsidence, that is, the second large period of pressure occurs on the working face.

4. Analysis on the Mechanism of Large and Small Period Breakage of Overburden in Stope

According to the mine pressure and rock layer control theory [16] and the results of similar simulation experiment, the overburden shape diagram in front of the working face is drawn, as shown in Figure 4.

Because of the effect of the support of the coal wall, the residual roof at the lateral collapse boundary of the goaf forms an inclined block. The rock stratum in the block bears the load of the overlying rock stratum and also transmits pressure to the low-level rock stratum. This roof with dual functions of bearing load and transferring stress is called the “inclined block bearing area”, as shown by the arrow in Figure 4.

When the caving horizon develops to the key stratum 1 at the lower level, several weak rocks above key stratum 1 collapse synchronously, forming a small periodic stope pressure (Figure 5a). After the small period of compression, the overlying strata controlled by key stratum 1 and the adjacent key stratum 2 are separated. The load borne by the key stratum 2 that does not collapse after the separation is transferred to both sides of the stope (in front of the coal wall and in the goaf), and the load in front of the coal wall is transferred downward through the inclined block bearing area, and acts on the stope. With the rise of caving horizon and the weakening of rock mass strength, the bearing area of inclined block develops in the direction of higher and wider, until the bearing area of inclined block reaches its maximum when key stratum 2 is fractured, forming large periodic pressure on the stope (Figure 5b).

Based on our literature review [12], when there is no confined aquifer above the thin bedrock stope, the topsoil load acting on the bedrock decreases significantly with the continuous progress of the stope. When there is a loosely confined aquifer, due to the mobility and recharge of confined water (as shown in Figure 3e), the load of the topsoil layer will be uniformly transferred to the lower bedrock through the aquifer. This is also proven by the stress test results during the experiment. At this time, the load undertaken by the bedrock is obviously greater than that of the unconfined aquifer, which provides an objective condition for the failure of the high key stratum.

As can be seen from Figure 5, with the continuous advance of the stope, the roof caving horizon gradually expands upward, and the bearing area of the inclined block expands. Before the collapse of key stratum 1, the bearing area of the inclined block is the smallest (zone ①). The fracture of key stratum 1 affects the synchronous fracture of several rock layers above it, causing strong disturbance to the roof above key stratum 1. Coupled with the double action of high stress transferred from loose confined aquifer to the bedrock surface, the primary cracks in the rock layer expand, secondary cracks develop, the transfer bearing capacity weakens and the range of the inclined block extends to the deep and upper part of the coal rock mass (zone ②). When the caving horizon further expands upward, the strength of rock mass is weakened, and the influence of caving angle on the transfer carrying capacity of the inclined block bearing area becomes apparent. The boundary rock mass above the caving line lacks the bearing foundation, and the transfer capacity is poor. The bearing area of the inclined block further develops towards the deep coal mass (zone ③).

5. Force Analysis of Hydraulic Support under Double Key Strata Condition

When there are two key strata in the overlying rock of the working face, and the distance between the two key strata is not large, the basic top of the stope (low key stratum 1) presents periodic fracture characteristics along with the mining of the working face. When the working face is drawn to a certain distance, the key stratum of the upper level and the key stratum of the position are broken at the same time, resulting in a complicated fracture phenomenon, and the action relation of the surrounding rock of the support is designed as shown in Figure 6.

It can be seen from Figure 6 that the formula for calculating the working resistance P₁ of the bracket under the condition of double key strata is as follows [17,18]:

(1)$P_1=(Q_{\mathrm{Z}}+Q_B)\cos \alpha \cdot K$

where: $Q_{\mathrm{Z}}=W\gamma L_Z$, $Q_B=Q_L+W\gamma \cdot h_2\cdot l+{\xi }_1{Q}_s$; $W$—the width of the bracket, m/frame; $\gamma$—the average bulk density of the immediate roof strata in the caving zone, kN/m³; $L_Z$—the length of direct roof beam, $L_Z=L_{\mathrm{d}}+L_h+L_{Zl}$ ($L_{\mathrm{d}}$—Face-to-tip distance, $L_h$—the sum of the length of the top beam and the front beam of the support, $L_{\mathrm{d}}+L_h$—the maximum roof control distance and $L_{Zl}$—the limit ceiling length of the immediate roof rock layer behind the support), m; $\alpha$—the coal seam dip angle, degrees; $K$—the safety factor, and the fully mechanized face is generally 1.2; $l$—the overall cut length of overburden, m; $Q_s$—the load transferred from the loose aquifer to the bedrock face.

As can be seen from Equation (1), the loads borne by the hydraulic support under the combination condition of double key strata include: weight $Q_Z$ of the direct roof, weight $Q_L$ of the broken block B of key stratum 1 and weight $W\gamma \cdot h_2\cdot l$ of the broken rock layer between key stratum 1 and key stratum 2. The load ${\xi }_1{Q}_s$ is transferred from the loosely confined aquifer and borne by the failure of the high key stratum to the support. The longer the fracture distance of the high key stratum is, the larger the ${\xi }_1{Q}_s$. $Q_s$ is the unique load of the thin bedrock working face, that is to say, the load borne by the hydraulic support of the working face of the thin bedrock coal seam is higher than that of the hydraulic support under the condition of no loose aquifer. When the composite fracture occurs in the high key stratum, the longer the fracture distance, the greater the load on the support and the greater the probability of a support crushing accident.

6. Practice of Safe Mining in Thin Bedrock Coal Seam Face

6.1. Technical Measures for Safe Mining of the Working Face

According to the rock mechanics test results, the uniaxial compressive strength of the rock sample of a high level rock stratum reaches 100 MPa, and the rock is hard and dense. According to our results (Section 4) and combined with previous working face mining experience, blasting forced roofing technology measures can be adopted for high key strata before working face mining [19,20] to artificially reduce the fracture distance of high strata, so as to reduce the probability of bracket pressing.

This forced caving adopts the deep hole loose blasting scheme, that is, in the semi-infinite medium, strong shock wave and a large amount of high-temperature and high-pressure explosive gas were generated after the explosive exploded in the gun hole. As the explosion pressure far exceeds the dynamic compressive strength of the medium, the medium in a certain range around the hole is strongly compressed and crushed to form a compression and crushing zone. In this area, a considerable part of the blasting energy is consumed in the excessive breakage of the medium, and then the shock wave is transmitted to the interior of the medium and propagates to the interior of the rock mass in the form of a stress wave. Under the action of the stress wave, the medium particle produces radial displacement, radial compression and tangential tension in the medium near the compression region.

6.2. Forced Caving Blasting Parameters of the Working Face

According to the previous research results, based on the engineering geological conditions of the 1611 (3) working face and the force bearing formula of the support (1), it is determined that the forced top setting distance of the high key stratum is 20 m. The hole layout is shown in Figure 7, and the blasting parameters are shown in Table 3.

The selected explosive single-roll explosive specification is φ63 mm × 1000 mm × 3.3 kg, detonated by an ordinary millisecond electric detonator. The blasting mud should be dried and screened, and the maximum particle diameter should be less than 8 mm. The explosive and containment structure are shown in Figure 8.

6.3. Effect Evaluation

Forced caving measures were taken on the 1611 (3) working face before mining to reduce the breaking distance of high key layers. The data of working resistance of hydraulic support monitored during mining were sorted out and generated as shown in Figure 9. The maximum working resistance of the hydraulic support is also within the range of rated working resistance. There is no support crushing accident, and the working face is safely stoped.

7. Conclusions

• (1). Based on the engineering geological conditions of thick loose bedrock coal seam working face 1611 (3), the physical simulation experiment of similar materials for thin bedrock coal seam mining is conducted, and it is concluded that the stope roof presents pressure characteristics of large and small periods in deep buried thin bedrock coal seam mining.

• (2). Under the action of load transfer of loose confined aquifers, caving horizons collapse successively from bottom to top. There is an inclined block bearing area above the stope of thin bedrock coal seam, which bears the load of the overlying strata and also transmits pressure to the lower strata. After a small period of weighing, under the double action of the fracture disturbance of the low key stratum and the load transferred from the aquifer to the bedrock, the primary cracks in the inclined block rock strata expand, secondary cracks develop and the transfer bearing capacity weakens. The range of the inclined block extends to the deep and upper part of the coal rock mass until the inclined block bearing area reaches its maximum when the high key stratum is fractured, forming the large periodic pressure of the stope.

• (3). When the combination of high and low rock strata is broken, the load borne by the hydraulic support is much greater than that of the hydraulic support without an aquifer, due to the load transfer of the loose aquifer.

• (4). For the mining area of 1611 (3) high key stratum, forced roof caving measures are taken, and the blasting parameters are determined. After the roof pre-cracking, no support crushing occurs on the working face, and the effect is good.

Author Contributions

Conceptualization, L.L. and S.Z.; methodology, H.Z.; software, S.Z.; validation, L.L., Y.X. and S.Z.; resources, H.Z.; data curation, L.L. and S.Z.; writing—original draft preparation, L.L.; writing—review and editing, H.Z. All authors have read and agreed to the published version of the manuscript.

Institutional Review Board Statement

Not applicable.

Informed Consent Statement

Not applicable.

Data Availability Statement

Not applicable.

Conflicts of Interest

The authors declare no conflict of interest.

Publisher’s Note: MDPI stays neutral with regard to jurisdictional claims in published maps and institutional affiliations.

Figures and Tables

Figure 1. Cross-section view of the 1611 (3) working face in the Zhangji Coal Mine.

Figure 2. Physical model. (a) Water sand belt simulation diagram. (b) Water pressure balance system.

Figure 3. Migration characteristics of the stope overburden. (a) Stoping 50 m. (b) Stoping 70 m. (c) Stoping 85 m. (d) Stoping 120 m. (e) Stoping 130 m.

Figure 4. Partition of the stope overburden.

Figure 5. Fracture characteristics of large and small periodic weighing (The inclined bearing block is divided by ①, ② and ③ in the diagram). (a) Small periodic weighting of stope; (b) Large periodic weighting of stope.

Figure 6. Relationship between support and surrounding rock in the case of two key strata.

Figure 7. Arrangement plan of the blast.

Figure 8. Schematic diagram of charging and sealing structure.

Figure 9. Working resistance distribution diagram of the hydraulic support on working face.

Table 1

Strata lithology.

No.	Their Thickness (m)	Depth of Burial (m)	Rock Name
15			Unconsolidated formation
14	2.1	392.3	Weathered middle sandstone
13	1.9	394.4	Weathered sandstone
12	3.7	396.3	Weathered mudstone
11	2.3	400.0	Mudstone
10	1.8	402.3	Fine-grained sandstone
9	2.7	404.1	Medium fine-grained sandstone
8	5.4	406.8	Medium fine-grained sandstone
7	1.6	412.2	Silty fine-grained sandstone
6	4.8	413.8	Sandy mudstone
5	2.9	418.6	Mudstone
4	3.2	421.5	Argillaceous sandstone
3	6.7	424.7	Fine-grained sandstone
2	4.6	431.4	Mudstone
1	6.4	436	13-1 coal

Table 2

Mechanical parameters of the investigated strata.

No.	Lithology	Elastic Modulus (GPa)	Cohesion (MPa)	Internal Friction Angle (Degree)	Poisson’s Ratio	Ratio Number (Sand: Lime: Gypsum)
1	Unconsolidated Formation	8	2	25	0.4	75:10:5
2	Wind oxidation zone	8	4	25	0.35	72:13:5
3	Middle fine-grained sandstone	15	6	33	0.3	6:0.6:0.4
4	Sandy mudstone	14	3.2	31	0.34	7:0.6:0.4
5	Fine-grained sandstone	30	8	40	0.25	5:0.6:0.4
6	Mudstone	12	2.8	30	0.32	7:0.7:0.3
7	13-1 coal	4	1.5	23	0.33	10:0.5:0.5
8	Floor rock formation	25	6	35	0.27	6:0.5:0.5

Table 3

Bore parameters.

BlastholeSerial No.	Blasthole Depth (m)	Dip Angle (Degree)	Horizontal Angle (Degree)	Blasthole Diameter (mm)	Charge Length (m)	Stemming Length (m)
1#	42	32	60	94	36	6
2#	30	41	45	94	24	6
3#	24	50	30	94	18	6
4#	21	45	0	94	15	6
5#	24	28	−30	94	18	6
6#	30	19	−45	94	24	6
7#	42	10	−60	94	36	6

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