Mechanical behaviors of coal measures and ground control technologies for China’s deep coal mines -A review

Hongpu Kang,Fuqiang Gao,Gang Xu,Huaiwei Ren

a CCTEG Coal Mining Research Institute,Beijing,China

b Coal Mining Branch,China Coal Research Institute,Beijing,China

c State Key Laboratory of Coal Mining and Clean Utilization,China Coal Research Institute,Beijing,China

Keywords: Deep underground coal mine Mechanical behavior Mining-induced stress Mining-induced fractures Ground control for roadways Ground control for working face

ABSTRACT This paper reviews the major achievements in terms of mechanical behaviors of coal measures,mining stress distribution characteristics and ground control in China’s deep underground coal mining.The three main aspects of this review are coal measure mechanics,mining disturbance mechanics,and rock support mechanics.Previous studies related to these three topics are reviewed,including the geomechanical properties of coal measures,distribution and evolution characteristics of mining-induced stresses,evolution characteristics of mining-induced structures,and principles and technologies of ground control in both deep roadways and longwall faces.A discussion is made to explain the structural and mechanical properties of coal measures in China’s deep coal mining practices,the types and distribution characteristics of in situ stresses in underground coal mines,and the distribution of mininginduced stress that forms under different geological and engineering conditions.The theory of pretensioned rock bolting has been proved to be suitable for ground control of deep underground coal roadways.The use of combined ground control technology(e.g.ground support,rock mass modification,and destressing) has been demonstrated to be an effective measure for rock control of deep roadways.The developed hydraulic shields for 1000 m deep ultra-long working face can effectively improve the stability of surrounding rocks and mining efficiency in the longwall face.The ground control challenges in deep underground coal mines in China are discussed,and further research is recommended in terms of theory and technology for ground control in deep roadways and longwall faces.

1.Introduction

China is the world’s largest coal producer and consumer.In 2021,China’s coal production reached 4.13 billion tons,accounting for more than 50% of global coal production.With the quick consumption of coal resources,the shallow resources in the central and eastern regions in China would be depleted.At present,China’s coal mines are extending to the depth at a rate of 8-25 m annually.More than 50 mines have been mined to a depth of larger than 1000 m,with the deepest reaching 1510 m.

The definitions of deep mining can be summarized in two categories (Xie et al.,2006,2012,2015;He,2014).The first one is defined according to the mining depth,and the other is defined based on the environment changes for the coal measures or the occurrence of rock disaster,such as high geostress,high ground temperature,high osmotic pressure,and strong mining disturbance.The phenomena such as the large deformation of roadways(drifts),high mining pressure,coal and gas protrusion,impact ground pressure and other dynamic hazards are the signals of the beginning of deep mining.In this paper,deep mining is defined by the ratio of rock stress to the strength of coal measures.Generally,the rock stress is determined according to the following criteria:

(1) 0-10 MPa(low-stress condition);

(2) 10-18 MPa(medium stress condition);

(3) 18-30 MPa(high-stress condition);and

(4) >30 MPa(ultra-high stress condition).

A coal mine can be considered to be a deep mining when the rock stress exceeds 18 MPa,which is greater than the uniaxial compressive strength (UCS) of coal seam itself in many mining areas and close to the strength of weak rocks.Ultra-deep mining can be defined when the rock stress exceeds 30 MPa,which is greater than the UCS of soft rocks such as mudstone and siltstone in some mines.Based on the mining depth(more than 700 m),ultradeep mining is assumed when the mining depth is more than 1200 m.

Ground controls in roadways and longwall faces are the hot topics in deep mining,as shown in Fig.1.Compared with shallow coal mines,deep coal mines suffer high in situ stress and strong mining disturbance,resulting in large deformation,long duration,and serious damage to the roadways (Adams and Rennison,2003;Hucke et al.,2006;Lubosik and Prusek,2010;Coggan et al.,2012;Schloemer et al.,2014;Lubosik et al.,2015;Kent et al.,2016).The mining pressure at the longwall faces is excessive,and the phenomena such as coal wall bulking,roof collapse and hydraulic shield damage are prominent (Alber et al.,2009;Swanson et al.,2016;Gao et al.,2021;Jangara and Ozturk,2021).

Tremendous efforts have been made to deal with ground control in deep coal mines (Clifford et al.,2001,2002;Arthur,2006;Hucke et al.,2006;Bowler et al.,2009;Kent et al.,2014;Kukutsch et al.,2019;Batugin et al.,2021;Ma?kowski et al.,2021;Wu et al.,2021).Germany’s ground control technology for deep coal mines is the pioneer.The maximum mining depth of German coal mines exceeded 1700 m,and tremendous experiences had been accumulated for deep mining and ground control(Shi et al.,2009,2013).In terms of roadway support,ground control technologies,including U-steel yieldable arches,high-strength rock bolts and cables,and combined support of rock bolts+U-steel sets+back-frame filling,have been successfully applied in Ruhr,Saar and Aachen mine districts and promoted to some coal mines in China.By increasing the length of the longwall faces (the average length more than 300 m),the centralization of production can be realized and the excavation workloads of longwall entries are reduced.

China started to address the problem of deep coal mining in the late 1980s (He and Qian,2010;Kang et al.,2018a;He,2021;Xie,2021).For example,an in situ geomechanical testing system for coal measures was developed(Kang et al.,2007a,2019),enhancing our understanding of the characteristics and evolution laws of deep mines.For ground control of deep roadways,the pre-tensioned rock bolting theory was proposed.Many high-strength pretensioned rock bolting technologies were then developed (Kang et al.,2007b):

(1) The combined ground support-modification-destressing technology for deep roadways (Kang et al.,2020,2021a,b);

Fig.1.Schematic diagram of ground control in underground coal mines.

Fig.2.Laboratory tested mechanical properties of coal specimens under varying confining pressures: (a) Young’s modulus;and (b) Compressive strength.

(2) The support technology for deep entry retained along gob side (Kang et al.,2010a;Zhang et al.,2015);

(3) The combined support technology of rock bolts+U-steel sets+back-frame filling for deep roadways(Jiang et al.,2020);and

(4) The concrete-filled steel tube support technology in deep soft rock roadways (Gao et al.,2010).

For the ground control of deep longwall faces,the fracturing and caving patterns of roof strata,the formation conditions of bearing structure,and identification of roof stability have also been noticed(Wang et al.,2019,2020a,2021a).The mechanisms of the interaction between the hydraulic shields and the surrounding rocks have been studied,and the corresponding mechanical models have been established(Wang and Pang,2015).The method for calculating the working resistance of hydraulic shields has also been proposed,and a series of high-tech hydraulic shields has been developed,which can basically meet the requirements of ground control in deep longwall faces (Wang et al.,2020b,2021b).

This paper reviews the main results obtained in the fields of stress distribution characteristics,mechanical behaviors of coal measures and ground control technologies in China’s deep coal mines.The results mainly consist of two parts: the research completed by the Coal Mining Branch (CMB) of the China Coal Research Institute (CCRI) (CMB-CCRI),and the latest result made under the National Key Research and Development Program“Strata control and intelligent mining technology for deep coal mines with depth of 1000 m”.The existing technical problems and development trends of ground control in China’s coal mine are discussed.

2.Geomechanical characteristics of coal measures in deep coal mines

The geomechanical properties of coal measures,particularly coal seams,immediate roofs,and floors,play a key role in the stability of excavations in terms of roadways and longwall faces.

2.1.Strength of coal measures

2.1.1.Strength of coal

Fig.3.Axial stress-axial strain behavior of coal specimens under varying confining pressure conditions (modified from Gao and Kang,2016).

Coal is a solid,combustible,organic geomaterial with a large number of pores and fissures at different scales.The pores and fissures are typically filled with methane.The physical and mechanical properties of coal thus significantly differ from other sedimentary rocks.For example,the bulk density of coal is generally 1200-1400 kg/m3,which is significantly less than that of sedimentary rocks (2000-2600 kg/m3).Common coal mechanical properties include the UCS,tensile strength,Young’s modulus,Poisson’s ratio,and residual strength,which can generally be determined by laboratory tests.The CMB-CCRI Rock Mechanics Laboratory has performed a large number of tests on coal specimens since 2003,including 837 unconfined compression tests and 900 indirect tension tests on coal specimens obtained from more than 100 coal mines across China.

The statistical result of a large number of test data shows that the minimum and maximum UCSs of the tested coal specimens are 1.37 MPa and 65.56 MPa,respectively;the mean and standard deviation are 17.28 MPa and 10.15 MPa,respectively.This indicates that the coal specimens in China are generally weaker than those from South Africa,USA,and India,with average strengths of 22.93 MPa,22.1 MPa,and 23.5 MPa,respectively(Buzzi et al.,2014;Mathey and van der Merwe,2015).

In deep mining,the coal surrounding a roadway may be subjected to high confining pressures.It is important to obtain the coal strength under high confining pressures for ground control and pillar design.Fig.2 shows the Young’s modulus and compressive strength obtained from laboratory tests on standard coal specimens under varying confining pressures.The strength of the coal with a UCS of 20 MPa can reach up to 100 MPa under a high confining pressure of around 25 MPa.It is interesting that the dispersions of Young’s modulus and compressive strength of the two coal specimens do not become smaller as the confining pressures increase.

In addition to the peak strength of coal (i.e.UCS),its residual strength is also an important factor for analyses of ground control.When the coal surrounding an excavation site fails,it can result in considerable plastic deformation,and the residual strength will be reached owing to its relatively low peak strength and high mininginduced stresses.This is particularly true for deep longwall entries,pillars,and longwall faces.A good understanding of the coal residual strength is therefore of great significance for ground control in deep coal mines.

Fig.3 shows the axial stress-axial strain curves of coal specimens subjected to compression tests under a range of confinement conditions.Both the peak and residual strengths are found to increase with the confining pressure,and the residual strength increases at a faster rate than the peak strength.The residual-to-peak strength ratio as a function of confining pressure is plotted in Fig.4.The results are further investigated by analyzing laboratory compression test data from the literature based on a variety of rock and field tests conducted on coal pillars.Gao and Kang (2016) suggested that the fracture intensity has a limit influence on the residual strength of jointed rock mass,regardless of the confining conditions.

The above-mentioned coal properties refer to as the mechanical properties of intact coal specimens obtained from the laboratory.In in situ tests,a rod with a probe attached to the top was inserted into the borehole.The probe with an indenter was connected to a highpressure pump with a pipe.Under the hydraulic pressure generated by the pump,the indenter was pushed against the borehole wall until failure,identified by a sudden increase in the displacement of the indenter and a maximum pressure value.The pressure was recorded and used to estimate the rock strength,according to the relationship between the pressure and rock UCS.This was established by a series of laboratory tests on different types of rocks,as shown in Table 1 and Fig.5.Compared with the laboratory test results of coal specimens,the following characteristics were observed:

(1) For most of the coal mines,the coal strength is generally less than 15 MPa,and it is only about 5 MPa for some coal mines.

(2) In the same coal seam,the coal strength fluctuates greatly at different depths due to the uneven distribution of coal seam properties,joints,fissures and other structures.

Fig.4.Residual-to-peak strength ratio(σr/σp)of coal under varying confining pressure(σ3) conditions (modified from Gao and Kang,2016).

Fig.5.In situ compressive strength of coal seam in Changcun coal mine,Yima coal mining area.

(3) Excavation leads to the failure of the coal near the surface of a roadway.The coal strength is significantly reduced or even completely lost.This phenomenon is particularly prominent in deep roadways,where the crushing range of the coal ribs of the roadway is relatively large and the reduction of coal strength is more significant.

(4) The stress of surrounding rocks gradually increases as it moves away from the roadway surface,resulting in a tendency for the coal strength to increase with depth.

(5) Since the points of penetration tests were distributed in different coal mines,regions and coal seams,no mature correlation was found between the compressive strength of coal seams and the burial depth.

The coal strengths obtained from laboratory tests and in situ borehole penetration tests only reflect the strength of intact coal or small-volume coal,and cannot represent the mechanical properties of coal mass.A direct way to assess the mechanical behavior of coal mass is to perform in situ tests on large coal blocks.For example,Bieniawski (1968a,b) used hydraulic jacks to conduct in situ uniaxial compression tests on a large cuboid coal specimen(1 m×1 m×2 m)cut from a pillar of an underground coal mine in South Africa.However,this kind of in situ tests was very timeconsuming and cannot be routinely performed.With the rapid development of computer hardware and numerical software,a novel numerical method called the synthetic rock mass (SRM)method has been developed to assess the mechanical behavior of rock mass (Pierce et al.,2007).The SRM approach uses a discrete fracture network superimposed on a bonded particle model to represent a jointed rock mass.Gao et al.(2014a) used the SRM method to evaluate the mechanical properties of coal mass and assess how the properties change with specimen size.The results showed a distinct scale effect where the UCS and Young’s modulus decrease with increasing coal specimen size and approach a representative value,as shown in Fig.6.The variations of UCS and Young’s modulus also decrease with increasing specimen size.

2.1.2.Rock strength

A marked feature of China’s coal measures is the wide distribution of different types of mudstones in the roof and floor,including sandy mudstone,carbonaceous mudstone,calcareous mudstone,muddy slate,muddy shale,muddy siltstone,and muddy sandstone.The UCSs of these rocks generally range between 10 MPa and 60 MPa(Meng and Su,2014;Liu et al.,2021;Wang and Gao,2022).Fig.7 shows the Young’s modulus and strength of two mudstones obtained from laboratory tests on standard specimens.The two mudstones are relatively weak with a deformation modulus of around 1.5 GPa and a UCS of 10-40 MPa.Both the Young’s modulus and strength increase with confining pressure.The strength of the mudstone from one coal mine (Jinjitan mine) in Shaanxi Province increases significantly faster than that of the mudstone from the other mine (e.g.Dadijing mine)in Inner Mongolia Autonomous Region,China.The latter exhibits a weak strength of around 40 MPa even under a confining pressure of 30 MPa.

Fig.6.Numerically predicted mechanical properties of coal mass for different specimen sizes and loading directions with respect to the applied axial loading (modified from Gao et al.,2014a): (a) Young’s modulus;and (b) UCS.The different color coding indicates the loading direction with respect to the butt cleat and face cleat.

Abundant clay minerals can lead to severe weathering after exposure and significantly reduce the rock’s integrity and strength(Kang,1993).This consequently enhances the rock dilatancy and rheology,which eventually leads to large deformation,destabilization,and damage of the surrounding rocks.Fig.8 shows that the UCS and Young’s modulus of mudstone from one coal mine in Shanxi Province decrease continuously with increasing water content (Liu,2019).The peak strength and Young’s modulus for mudstone with 1.06% water were respectively 56% and 71% lower than those of dry specimens.The residual strength of mudstone also generally decreases with increasing water content,whereas the Poisson’s ratio gradually increases.This indicates increased lateral deformation of the rock specimens.

In situ strength tests using the penetration method were performed to evaluate the strength of roof rocks in deep coal mines.Some of the results are listed in Table 2.

According to Table 2,the following findings can be obtained:

(1) There were distinct fluctuations in the strength at different depths of the same roof rock,mainly due to the uneven distribution of rock structures and discontinuities.

(2) The roof near the surface of the roadway was damaged and its strength was significantly reduced.At some positions,the roof was so broken that its strength could not be measured.

(3) Similar to the coal strength,the strength of roof rock tended to increase with depth as it moved away from the roadway surface.

2.2.In situ stress field in deep coal mines

In situ stress measurements in coal mines are the most direct and effective method to assess pre-mining stresses in coal measures.The overcoring method and hydraulic fracturing (HF)method have been widely used in coal mines(Kang et al.,2009).In situ stress measurements are generally performed in competent and unfractured rocks to meet the assumption of ideal rock behavior (i.e.continuous,homogeneous,isotropic,and linearelastic).However,most coal seams are relatively soft and highly fractured compared with rocks,which limits the pre-mining stress data measured in coal seams availably.Since 2000,the CMB-CCRI has conducted in situ stress measurements in more than 40 coal mining areas using in situ stress measurement device by HF with small-diameter borehole (Kang et al.,2007a,2010b).Combined with field stress measurements in coal mines(Zhou et al.,2005;Cai et al.,2006;Wang et al.,2010;Zhang et al.,2012),a total of 1357 data were used to establish a database of in situ stress in coal mines in China(Kang et al.,2009).Fig.9 shows the ratio of the maximum(KH),minimum(Kh),and average(Kav)ratios of the horizontal to the vertical principal stress,and the following expressions were obtained by regression analysis:

whereHis the depth.

Table 3 lists the measured in situ stresses at typical deep coal mines,where σHand σhare the maximum and minimum horizontal principal stresses,respectively,σVis the vertical stress,and Δσ is the difference between the maximum and minimum principal stresses.The depth of the in situ stress measurement points ranged from 718 m to 1342 m,representing different depths of deep coal mines.Compared with shallow coal mines,the in situ stress distribution of deep mines has the following characteristics:

Fig.7.Laboratory tested mechanical properties of two mudstones under varying confining pressures: (a) Young’s modulus;and (b) Compressive strength.

Fig.8.Axial stress-axial strain curves of mudstone with varying water content(modified from Liu,2019).

Table 2 In situ compressive strength of muddy roof rocks in typical deep coal mines.

(1) The in situ stresses in shallow mines are relatively small,and all three principal stresses tend to increase with depth.There are marked dispersions in the maximum and minimum horizontal principal stresses,and the dispersions in shallow mines are greater than those in deep mines.The number of the coal mines with the vertical stresses greater than 19 MPa are 19 out of 20 measurements in Table 3,and the smallest one is close to 18 MPa.For the measurements in the mines with depth greater than 1000 m,the maximum horizontal stress all exceed 30 MPa,and the maximum reaches 42.19 MPa.

(2) The reverse fault stress regime plays a dominant role in shallow coal mines,where σH>σh>σV.For the eight deep mines listed in Table 3,two types of in situ stress regimes exist,i.e.strike-slip fault stress regime (σH>σV>σh) and normal fault stress regime(σV>σH>σh),which accounts for 45% and 55%,respectively.In the Huainan,Yima and Pingdingshan coal areas,the vertical stress is greater than the maximum horizontal principal stress at most of the testing points,and the stress regime is of a typical normal fault regime.However,the strike-slip fault regime is dominant in Xinwen coal mining area with significant tectonic stress.

(3) The difference between the maximum and minimum principal stresses is 9.54-19.39 MPa,mainly in the range of 11-18 MPa.The difference in the principal stresses represents the shear stress of the rock,which is relatively large in deep mines.

(4) For the ratio of horizontal principal stress to vertical stress in shallow mines,the dispersion of this ratio gradually decreases with increasing depth and tended toward a certain value.In Table 3,the range ofKavfor deep mines is 0.55-1.13,and higherKavvalues mean that the harder and intact rocks are more influenced by the geological structure.For deep mines with depth greater than 1000 m(measurements 1-10 in Table 3),Kavis generally 0.74-1.13,with a much narrower distribution.

(5) In Table 3,the ratio of the maximum horizontal principal stress to the vertical stressKHis 0.71-1.49,and the measurements in the range of 0.8-1.2 accounts for 75%,indicating that the difference between the maximum horizontal principal stress and the vertical stress in deep mines tends to become insignificant.The ratio of the minimum horizontal principal stress to the vertical stressKhis 0.39-0.77,mainly distributed in the range of 0.44-0.77.This is different fromKHthat has a large difference,and the distribution of horizontal stress in deep mines had an obvious directionality.

Fig.9.Ratio of the horizontal principal stress to the vertical stress as a function of depth.Data were collected from in situ stress measurements in typical underground coal mines in China (modified from Kang et al.,2009).

The distribution of the ratio of rock strength to in situ stress is listed in Table 4,where σcis the rock UCS.The ratio of rock strength to in situ stress is low.For mudstone,most of the ratios(compressive strength to the maximum principal stress) are less than 2,some close to 1,and the minimum value is only 0.87.Therefore,for deep mines,it is likely that the surrounding rock mass will be damaged after excavation of the roadways.

Table 3 In situ stress measurements in typical deep coal mines.

Table 4 The ratio of muddy roof rock strength to in situ stress in typical deep coal mines.

2.3.Structural characteristics of coal measures

The structure of coal measures has some unique features,in addition to the characteristics of general sedimentary rocks.Coal seams contain a large number of pores and fissures with scales ranging from macroscopic faults,laminae,and joints to micronand nano-sized pores.The internal structure of coal seams significantly affects the physical and mechanical properties.

Fig.10 shows the distribution of geological structures in the Huainan coal district (Yuan,2006),a typical coal area with deep mining and soft rocks.In this area,normal faults,reverse faults,synclines,and anticlines are densely distributed.These large-scale structures have led to remarkable disturbances in the thickness and distribution of the coal seams,and the tectonic stresses have complicated the distribution of in situ stress.Small-scale laminations including joints and other structural planes in the roof and floor of coal seams compromise the stability of the surrounding rocks of underground openings.

Coal fissures are can be classified as endogenous or exogenous fissures according to their genesis (Su et al.,2002).Endogenous fractures are cracks that form during the coalification process,and the most typical fractures are cleats,which can be further divided into face cleats and butt cleats.Face cleats generally form first,and butt cleats are mutually orthogonal and perpendicular to bedding planes.Butt cleats generally terminate at face cleats,as shown in Fig.11.Exogenous fissures form in coal seams under tectonic loading and their scale varies significantly.For example,some fissures can extend into the roof and floor of the coal seam.The existence of these fissures and pores leads to significant differences in the physical and mechanical properties from those of rocks.

3.Distribution and evolution characteristics of mininginduced stresses in deep coal mines

Rocks surrounding underground openings in coal mines suffer in situ stresses and mining-induced stresses caused by extraction of the longwall panels.The latter can reach up to 2-5 times the in situ stress,which is the most significant feature of coal mines compared to other underground engineering.

The longwall mining method is the most commonly used in China’s coal mines.The distribution of mining-induced stress is related to many factors including the layout of longwall panels and entries,mining parameters of longwall face,spatial and temporal relationship with other longwall faces and excavations,and geomechanical parameters of coal measures.There are three main systems of longwall panel layout (Fig.12): single entry system,double or multiple entry system,and double-use entry system.For deep mining,double or multiple entry system is not generally used due to the difficulty in maintenance.Most of the deep mines in China adopt the single entry system,and longwall panels are separated by coal pillars with a certain width.In a double-use entry system,the headgate of the previous longwall panel is used as the tailgate of the current longwall panel,saving the coal pillars between the two longwall panels,which can improve the coal resource recovery rate.This method has advantages in deep mining due to less excavation and maintenance of entries.The double-use entry system is commonly used in Germany.The length of the longwall face is a very important mining parameter.In China’s coal mines,the length of a longwall face is generally 200-350 m,with the maximum length up to 450 m.For deep mining,long and extralong longwall faces can reduce the total length of the longwall entries.In Germany,deep mines commonly adopt 350-500 m extra-long longwall faces.

3.1.Mining-induced stress of single longwall panel

For single longwall panel where no other excavations are involved,when the longwall face proceeds,the immediate roof collapses and caves into the mined-out area,and the disturbed roof strata gradually extend upwards.It has been well documented thatthree distinct zones(i.e.caved zone,fractured zone,and continuous deformation zone) exist in the overburden rock strata above the mined-out area of the longwall panel.Evaluation of the three distinct zones has a significant influence on the distribution of mining-induced stress.Knowledge of the stress change is very important for understanding the caving process,and designing both face and roadway support.

Numerical simulation is an effective means to evaluate the mining-induced stress under a given condition.Fig.13 shows the vertical stress distribution simulated around a deep longwall panel as the longwall face advances,where the universal distinct element code(UDEC)was used for simulation.The extraction of coal results in stress concentration in the un-mined coal and the roof adjacent to the opening.When the vertical stress in the immediate roof and floor is relieved,the areas of both the stress concentration zone and stress relief zone increase with face advance.When the face has advanced a long distance(e.g.72 m),the immediate roof collapses and caves,resulting in closure of the mined-out area.As the longwall face continues to advance,the overburden pressures are transferred to the un-mined zone around the mined-out area(i.e.the front and rear of the longwall panel in this two-dimensional(2D)model).

Fig.10.Distribution of geological structures in the Huainan coal district (modified from Yuan,2006).1 -Normal fault;2 -Reverse fault;3 -Anticline;4 -Syncline;5 -Lower Paleozoic strata;6 -Lower Paleozoic and Taikoo strata.

Fig.11.Face cleat and butt cleat in a coal seam.

The stress changes can be analyzed by plotting the vertical stress at selected monitoring points with face advance.The vertical stresses monitored at 12 points in the immediate roof are shown in Fig.14.These 12 points can be considered to represent three different types of mechanical behaviors according to the observed patterns of stress change.The first type involves P1 and P12,which are located in the immediate roof above the un-mined coal.At P1,the simulated vertical stress increases gradually as the face advances away and reaches the peak of 60 MPa when the distance to the face exceeds 84 m.When the vertical stress is larger than this value,no obvious change in the stress is observed with further face advance,indicating an eventual stress equilibrium condition at P1.At P12,the vertical stress starts to increase when the face is approximately 84 m away from the monitoring point.These stress change patterns,at P1 and P12,indicate that mining-induced stress changes can be up to 80-90 m ahead of the modeled longwall panel.The second type involves P5 and P8,which are located in the central part of the immediate roof.At these two points,the vertical stress change is characterized by three distinct stages:a significant stress increase as the face approaches,a sharp stress drop to zero as the face passes,and a gradual increase as the roof caves and compacts.The third type involves P2-P4,P6,P7 and P9-P11.At these points,the vertical stress initially increases as the face approaches and then sharply drops to zero after the face passes.At P11,the vertical stress starts to increase when the face is approximately 84 m away,confirming that the influence area of the mininginduced stress changes can reach up to 80-90 m in front of the face.It should be noted that the monitored peak vertical stress increases from 35.5 MPa at P3 to 61.6 MPa at P11,indicating that the front abutment pressure increases as the face advances.

3.2.Mining-induced stress around longwall panels and pillars

When there are other longwall faces or gobs around a working face,the mining stresses will affect each other and superimpose,as shown in Fig.15.A certain abutment pressure is generated in front of and to the right of the gob;while near the gob,it is superimposed with the residual abutment pressure of the previous gob,resulting in a very high abutment pressure near the left corner of the longwall face.A high stress concentration is noticed on the coal pillar,making the maintenance of the roadway more difficult.

FLAC3D (fast Lagrangian analysis of continua in 3 dimensions)was used to simulate the distribution of mining stress at the extralong longwall face of the Kouzidong coal mine of Xinji Energy Company Limited,China.The 121304 longwall face had a coal seam thickness of 5 m,a depth of 1000 m,and a longwall face length of 350 m.The three-dimensional (3D) stress distribution around the longwall face is shown in Fig.16,and the distribution of abutment pressure near the gob side is shown in Fig.17.

The dynamic migration of mining-induced stress occurred along the proceeding direction of the longwall face.At the initial advancement,the front peak abutment pressure and its influence range were small,and there was no recovery of stress in the gob.As the face advanced and the mining area increased,the front peak abutment pressure and its influence range increased gradually.The stress in the gob area recovered,and the degree of recovery increased continuously.When the advancement of the longwall face increased to 180 m,the development of mining-induced stress field was basically in a stable state.

Stress measurements are an important method to understand the distribution of mining-induced stress.Four borehole stress gauges were installed at 120 m in front of the longwall face in the tailgate to measure the increases in the front abutment pressure of the 121304 longwall face.The monitoring results are shown in Fig.18.The front abutment pressure was observed around 100 m ahead of the longwall face,which increased gradually as the distance from the longwall face decreased.The front peak abutment pressures occurred at 10-12 m ahead of the longwall face.Beyond this,the coal was damaged and the bearing capacity was reduced,and the abutment pressure decreased continuously,reaching the minimum value near the coal rib of the longwall face,which was approximately equal to the coal residual strength.It can be seen that the numerical results of front abutment pressure of the 121304 longwall face was consistent with the measured data.

3.3.Mining-induced stress around entry retained along gob

Construction of entry retained along gob can be divided into six stages according to the mining-induced stress it suffers from excavation to scrapping.The six stages are excavation stage,stabilization stage,stage affected by the front abutment pressure of the first longwall panel,stage affected by the rear abutment pressure of the first longwall panel,stabilization stage with the rear mining influence,and stage affected by the front abutment pressure of the second longwall panel.Compared with the normal roadway,the stress,deformation and damage of the rocks surrounding the entry retained along gob have the following characteristics:

(1) Temporal characteristics of mining-induced stress distribution.The entry retained along gob suffers mining-induced stress from two longwall panels.The duration of this process is much longer than that of normal entries.In front of the first longwall face,the abutment pressure distribution is similar to that of normal longwall faces.However,in the back of the longwall face,one side wall of the entry no longer exists,causing the roof to fail and collapse.The abutment pressure changes dramatically,and the disturbance causes the roadway to take a longer time to stabilize.This phenomenon is frequently reported in deep roadways.As the second longwall face advances,the front abutment pressure is superimposed on the disturbance caused by mining the first longwall panel,leading to severe failure of the roadway.

(2) Spatial characteristics of mining-induced stress distribution.The entry retained along gob suffers mining-induced stress from two longwall panels.The stress in the surrounding rocks along the entry has a close relationship with the position (front or back of the longwall face) and distance.At a certain distance in front of the first longwall face,the surrounding rocks are affected by mining (the deep longwall face can be more than 100 m),and the front abutment pressure reaches its maximum at a certain distance from the longwall face.In the back of the longwall face,because the coal seam has been mined along one side of the entry retained along gob and a side packing is placed,the rear abutment pressure changes significantly and stabilizes at a certain distance.For a deep longwall face,the distance can be 70-150 m and more than 200 m in some cases.The entry retained along gob suffers mining-induced stress again at a certain distance ahead of the second longwall face,and the distance generally exceeds the first longwall face.

Fig.12.Typical longwall panel layout systems in China’s coal mines.

(3) The relationship between the depth and entry retained along gob.Many factors influence the stress and deformation of the surrounding rocks along an entry retained along gob,including the geological factors (e.g.the thickness,strength and stability of the coal seam,the distribution,strength and stability of the roof,the depth of the entry,and the distribution of rock stress) and engineering factors (e.g.the mining height,the profile and size of the entry section,and the entry support pattern and parameters).For deep mining,the depth is an important factor.In contrast to other panel layout systems with a chain pillar between two adjacent longwall panels,field experiences show that the depth has a much smaller impact on the deformation of the entry due to the structural characteristics of the rock mass surrounding the entry retained along gob.It is more favorable to the promotion and application of the double-use entry system in deep coal mines.

4.Rock structure and its evolution around deep underground openings

According to the mining disturbance,roadways can be divided into two classes,i.e.the ones affected by the mining of longwall faces and the ones not.The roadways affected by the mining account for more than 80%of the total roadways.A deep roadway has the characteristics of high geostress and strong mining disturbance.Weak,soft rocks surrounding underground openings (roadways and longwall faces) tend to exhibit large deformation,and brittle,competent rocks tend to experience burst failure.

4.1.Rock structure and failure characteristics of deep roadways

This section reviews the failure characteristics of deep roadways under different geological and engineering conditions,including the main roadways with soft rock and high rock stress,longwall entries with depth of 1000 m,deep entry retained along gob,and coal burst-prone entries.

4.1.1.Main roadway with soft rock and high stress

The main roadways in coal mines are generally excavated in strong rock layers.In deep conditions,main roadways are likely to experience severe deformation due to release of high in situ stress.This is more pronounced when the roadway is excavated in moisture-sensitive soft rocks such as mudstones.

The Kouzidong mine is a typical deep coal mine in China,with a depth of 1000 m.The roof and floor of main roadways in this coal mine are mainly composed of soft mudstones,with a mean UCS of 25.1 MPa (Table 2).A variety of support technologies have been applied to the soft rock roadways,including rock bolts and cables+shotcrete,rock bolts and cables+shotcrete +36U steel sets,rock bolts and cables+shotcrete +36U steel sets+antibottom arches and steel pipes filled with concrete,all of which cannot effectively control the large deformation of roadways.Rehabilitation had to be repeated to maintain the roadway function.

Fig.13.Simulated vertical stress distribution with longwall face advance (modified from Gao et al.,2014b).

Fig.14.Simulated vertical stress changes in the immediate roof above the gob.The 12 monitoring points were located in the immediate roof.Different intervals of horizontal axis represent the numerical calculation steps performed during the 10 mining stages which are sufficient to bring the model to equilibrium after excavation at each mining stage (Gao et al.,2014b).

Fig.19 shows the typical damage characteristics of main roadways with soft surrounding rock mass and high rock stress in the Kouzidong coal mine.The roof mudstone has low strength and can be easily weathered when exposed to water.Excavation-induced fractures expand gradually from shallow to deep,leading to a significant disintegration of the surrounding rocks.In the roof,rock mass bulking took place between rock bolts and cables (Fig.19a),and bulking failure was severe and had a marked time effect.The bulking depth reached 600 mm after 2-3 months and kept accumulating,which deteriorated the roadway stability.The roadway ribs showed a distinct step-bulge deformation characteristic which occurred at 0.5 m from the roadway floor where the support at the roadway bottom corner was relatively weak,and the maximum convergence reached up to 1.2 m(Fig.19b).The deformation of the roadway floor had a close relation with the structural rheology.The measured floor heave reached 1.5 m,accounting for about 80% of the roof-to-floor convergence(Fig.19c).In some localized areas,the accumulated floor heave reached up to 5 m,which was equivalent to the initial height of the roadway.Severe floor heave and wall-towall convergence were the main characteristics of these main roadways excavated in soft mudstones (Fig.19d).

4.1.2.Longwall entries with depth of 1000 m

Longwall entries are generally excavated in coal seams which are relatively weak compared to rock layers.Longwall entries suffer high in situ stress and mining-induced stress,and thus are more vulnerable to severe deformation.Taking the 121302 headgate of the Kouzidong coal mine as an example,the depth of the longwall entry was 1000 m,and the roof and floor were mainly of mudstone and sandy mudstone with high clayey mineral content.The mudstones were easily weathered and softened in contact with air and water.The entry section had a horseshoe-shaped cross-section with a width of 5.8 m and a height of 4.1 m.It was driven into the roof,leaving 2-m thick coal at the floor.A combined support system of rock bolts and cables+shotcrete+grouting was used for the ground control of the entry.Under high geostress,the entry exhibited a distinct time-dependent deformation,characterized by massive bulking failure of coal pillar and severe floor heave (see Fig.20).During the longwall mining period,a cumulative floor heave of more than 6 m was removed to keep the entry open for transportation and ventilation,and the entry side walls had to be trimmed around 2 m and re-supported to maintain the entry’s profile.

Field observations suggested that the side wall movement was greater on the pillar side than that on the working face side,exhibiting a spatial asymmetric large deformation.Severe rock mass bulking caused failure of some bolts and bulging of mesh and straps along the entire entry.Due to the high geostress and strong mining disturbance,sudden breakage of cantilever above the coal pillar increased stress concentration of the coal pillar and reduced the integrity and bearing capacity of the surrounding rocks.The broken zone in the shallow part of the entry extended to a greater depth,which further aggravated the dilation of the entry.

To examine the failure of entry,bore televiewer was used at the two sides of the entry.At the side of working face,a large range of loose and broken coals was observed in the range of 0-1 m and 2.1-3.2 m.A local cavity area existed in the range of 3.7-3.9 m and 4.1-4.3 m,and a localized broken area existed in the range of 5.9-6.4 m.The drilling depth of the coal pillar side was 6 m,but the collapse of the hole occurred at the depth of 3.5 m.To obtain the damage range of the coal pillar side,borehole core logging was performed.In Fig.21,coals in the range of 5 m were extremely broken,which caused two major problems.One was that the broken coal with poor integrity and low strength had insufficient bearing capacity.The other was that the rock bolts and cables had low bonding strength in the broken coal pillar,resulting in aggravation of the entry deformation.

4.1.3.Deep entry retained along gob

In deep coal mines,high in situ stress and strong mining disturbances could lead to distinct behaviors of entry retained along gob,such as extensive roof sag,severe floor heave and wall-to-wall convergence.Fig.22 displays the 5121B10 tailgate at the Xieyi coal mine,Huainan coal mining area.The average thickness of the coal seam was 1.4 m,with an inclination angle of 22°.The depth of the entry was about 700 m.The trapezoidal cross-section was inverted,with a middle height of 2.8 m and a width of 5 m.The maximum convergence at the high side was 1.5 m.The roof bolt at the corner of the high side had sunk into the coal,and the uppermost rib bolt was on the same horizon as the roof due to roof sag.The roof was seriously broken with bulging,and the steel strap was curved to the interior of the entry.Convergence was also observed at the low side,which was smaller than that at the high side.Serious floor heave took place in the entire entry.

Fig.15.Schematic diagram of the distribution of mining-induced stress.

Fig.16.Numerically simulated distribution of mining-induced stress around the two longwall panels at Kouzidong coal mine.

Fig.17.Numerically simulated abutment pressure at the 121304 longwall panel at Kouzidong coal mine.The curves are generated by plotting the vertical stress along a horizontal line parallel to the mining direction.Different colors refer to different mining distances.

4.1.4.Coal burst-prone entry

Fig.18.Field measured increases of front abutment pressure at the 121304 longwall panel at Kouzidong coal mine (Wang et al.,2020b).

Coal burst is a very prominent dynamic hazard encountered in deep coal mines(Mark,2016).As the mining depth increases,coal burst becomes more severe.Coal burst mainly occurs in roadways,especially longwall entries suffering from mining disturbances.The frequent occurrence of large energy events can lead to the breakage of the roof bolts and cables,extreme wall-to-wall convergence,and severe floor heave.Fig.23 shows the deformation and damage of the 21220 tailgate in the Changchun coal mine,Yima coal mining area.The maximum mining depth of the tailgate was 815 m.The tailgate was supported by a three-stage support system,i.e.rock bolts and cables+36U steel arches+hydraulic lifting sheds or portal brackets.Even with this competent support system used,large deformation still occurred after multiple high-energy seismic events (i.e.energy greater than 105J).The maximum wall-to-wall convergence reached up to 3 m,and the maximum floor heave and roof sag reached up to 1.5 m and 1 m,respectively.The 36U steel arches deformed and failed at many locations,and the steel arches also failed due to large torsional deformation in some parts.A large number of roof bolts and cables were buried by the mesh pocket induced by the broken surrounding rocks.The whole entry was rehabilitated 1.5 times that of the average repairs during the service period.In other words,the three-stage composite support system was not effective,as the supports did not function well and individual components failed.The subsequent secondary and tertiary supports were also damaged due to bending and twisting deformation,resulting in severe deformation of surrounding rocks.

4.2.Stratum structures and failure characteristics above deep longwall faces

4.2.1.Roof caving and periodic weighting

As longwall mining moves along the mining direction,the roof strata above the mined-out area may collapse periodically,which is a unique phenomenon associated with longwall coal mining.Mining disturbance and mining-induced stress are closely related to this phenomenon.Fig.24 shows the caving process of a physical model as the longwall face advances.During initial mining advancement,the immediate roof remains stable and the minedout area is left open.As the longwall face moves further,the immediate roof begins to bend and sag and eventually separate from the main roof along the bedding planes.The initial caving happens at mining step 9.At mining step 12,the second caving occurs.The caved zone involves three rock layers overlying the immediate roof,resulting in a height of around 18 m of the caved zone.As the longwall face continues to advance,the immediate and main roofs become a cantilever beam that collapses when its length reaches around 20 m at mining step 16.The fractured zone extends 33 m upward,reaching 51 m into the overlying strata.As the longwall face continues to advance,the periodic collapse takes place several times at mining steps 22,28,31,36,and 37,respectively.The height of the fractured zone increases at every periodic collapse.The periodic weighting distance varies between 10 m and 30 m.At each periodic collapse,the collapse zone involves not only the immediate and main roofs that compose the cantilever beam,but also the strata right above the fractured zone resulting from previous caving.This suggests that the periodic weighting is related to not only the local immediate and main roofs that compose the cantilever beam,but also a large global structure across the entire mined-out area.Dealing with the immediate roof alone may not obtain a substantial release of the front abutment pressure in front of the longwall face.It can be anticipated that this phenomenon will no longer exist when the failure of overlying strata has extended up to the ground.

Fig.19.Severe deformation of a main roadway with soft surrounding rocks and high rock stress in the Kouzidong coal mine:(a)Rock mass bulking between rock bolts in the roof;(b) Step-bulge deformation at the side wall;(c) Severe floor heave;and (d) Overall extreme squeezing failure of the roadway (modified from Jiang et al.,2020).

The physical model also shows a distinct periodic pattern of roof bridge failure of the main roof strata while the longwall mining continues to advance.The rupture of a rock bridge occurs periodically alongside periodic weighting as the longwall face advances,as shown in Fig.25.Periodic weighting can be detected by monitoring the working pressure of the hydraulic shields at the longwall face.The rupture of a rock bridge in the overburden strata behind a mined-out area is generally detected as seismic shakes.The energy released during rock bridge rupturing in the form of seismic waves may trigger a dynamic failure of underground openings (e.g.rock bursts,and coal bursts).

Numerical study was conducted to evaluate the factors that influenced roof bridges using a comprehensive finite element and discrete element software tool Elfen by changing one factor while the others remain constant.Two factors were evaluated: (1) the stress regime,which is represented by the horizontal-to -vertical stress ratio;and(2)the competence of the main roof layers,which is characterized by the stiffness and strength.Fig.26 shows the effects of the two factors on the roof bridge length.All curves of roof bridge length versus longwall mining distance show a similar pattern,which is characterized by an initial increase until the maximum value is reached,followed by a decrease,regardless of the geological and geotechnical conditions.In contrast to the hydrostatic stress regime,a higher or lower stress ratio tends to promote the formation of a longer roof bridge.For a higher stress ratio,the high horizontal stress provides higher lateral containment to the roof strata,which results in a longer span of the roof strata.For a lower stress ratio(<1),less fracturing occurs under the low horizontal stress.For the case of a stress ratio of 1.5,the initial roof collapse occurs when the longwall advances 100 m at a mining step of 20,which is considerably longer than that determined in the base case when the longwall advances 60 m.The strength and stiffness of the main roof show a strong influence on the collapse pattern of overburden roof strata.A competent main roof can thus sustain higher ground pressure and maintain intact for a longer time,leading to a longer initial collapse distance and longer roof bridge.

4.2.2.Deep extra-long longwall faces

Fig.20.Failure pattern of deep longwall entry suffering high mining-induced stress in the Kouzidong coal mine:(a)Severe bulking of coal pillar that caused the failure of some bolts and bulging of mesh and straps;and (b) Extensive floor heave.

Fig.21.Borehole core logs at the coal pillar side of a deep longwall entry in the Kouzidong coal mine.

The evolution of mining-induced stress in extra-long longwall faces with depth of 1000 m is distinct from normal longwall faces in terms of both the advancing and vertical directions.The extralong longwall face increases the probability of generating largescale fractures within the main roof above the gob area,forming zonal breakage characteristics,as shown in Fig.27.The roof initially breaks in the middle of the longwall face and then undergoes zonal breakage along the two longwall entries.The zonal breakage of the main roof leads to a non-uniform spatial distribution of the fractured rock blocks,namely,small blocks in the middle of the longwall face and large blocks at both ends.The characteristics of main roof breakage lead to the distribution characteristics that include lower pressure in the middle of the hydraulic shields and higher pressure at the two ends,which differ from that for conventional longwall faces,namely,higher pressure in the middle and lower pressure at the two ends,i.e.“O-X”pattern (Wang et al.,2019).

Field measurements in the Kouzidong coal mines showed that the distribution characteristics of the hydraulic shield resistance pressure along the longwall face differ from those obtained in the longwall faces with conventional lengths and medium mining depths.The pressures at both ends of the longwall face were found to be greater than those in the middle of the longwall face(Fig.28).This was attributed to the zonal breakage pattern that causes the breaking degree of the roof at both ends to be greater than that of the middle part of the longwall face.The roof fragments at both ends were found to be larger than those in the middle of the longwall face.The relatively small roof fragments in the middle of the longwall face resulted in more articulated contacts between the fragments,which increased the degree of freedom of the articulated system.According to structural mechanics theory,higher degree of freedom in an articulated system is indicative of higher instability.The smaller rock fragments in the roof of the middle longwall face apply lower pressure on the hydraulic shield than those on the roof at the two ends.

Fig.22.Deformation of an entry retained along gob at depth of 700 m in the Xieyi coal mine: (a)Irregular profile of the entry after deformation(the original profile was inverted trapezoidal);and (b) Extensive convergence at higher side wall.

Fig.23.Deformation of a deep entry suffering multiple seismic activities in the Changchun coal mine:(a)Severe closure of the roadway profile;and(b)Breakage of the steel arch(modified from Kang et al.,2015a).

Fig.24.Caving process captured by the physical model as the longwall face advances.The colorful lines indicate the boundaries of the fractured zone (modified from Gao et al.,2022).

Fig.25.Periodic collapse and formation of a roof bridge simulated by a physical model of longwall panel excavation: (a) Roof bridge formed before a periodic collapse of the cantilever beam;and(b)A new roof bridge formed after the collapse of the cantilever beam (modified from Gao et al.,2022).

5.Theories and technologies of ground control in deep underground coal mines

After years of research and practice,China has developed many ground control technologies for coal mine roadways,which can be divided into five types,as shown in Fig.29 according to the principles of ground control.The surrounding rock mass control technologies can be divided into the following 5 forms:

(1) Surface support.A constraining force is applied on the roadway surface to control the deformation of the surrounding rocks,including various kinds of framed supports and pillars,shotcrete,and poured concrete support.

(2) Anchoring of roadway rock.The support elements not only act on the roadway surface but also penetrate deep into the surrounding rock mass,including different types of rock bolts and cable bolts.

(3) Reinforcement.The mechanical properties of the surrounding rocks are improved by injecting grout into the fractured rock masses.

(4) Rock stress relief.The roadway stress state is improved by reducing or transferring high stress by various artificial pressure relief techniques.

(5) Combined support system.A combination of two or more methods is used to constrain the roadway deformation.

Deep roadways have the characteristics of high geostress,extensive mining disturbance,and large time-dependent deformation or burst failure.According to the characteristics of deformation and damage of deep roadways,the theory and technology system of ground control of roadways in deep coal mines are proposed.The system involves adoption of high-strength pretensioned rock bolts and cables as basic support,working in combination with the modification of rock masses and distressing,as shown in Fig.30.

5.1.Pre-tensioned rock bolts and cables

Rock bolting is the main support method of coal mine roadways,accounting for more than 75%.For deep roadways,rock bolts and cables are the preferred and basic support method when the anchorage force can meet the design requirements.Aiming at the deformation characteristics of the surrounding rocks of deep roadways and the role of rock bolt support,the theory of pretensioned rock bolt support for deep roadways is proposed,as shown in Fig.31.The main points of the theory are summarized as follows:

(1) The essence of rock bolting is to control the discontinuous and uncoordinated dilatant deformation of surrounding rocks (e.g.tensile and shear cracking,separation and sliding of discontinuities,and rock block rotation)in order to reduce the strength deterioration of surrounding rocks and maintain its integrity and self-supporting capacity.

Fig.26.Factors affecting the roof bridge length: (a) Horizontal-to-vertical stress ratio;and (b) Main roof competence (modified from Gao et al.,2022).

Fig.27.Schematic diagram of the zonal breakage of roof layers above a mined-out area of an extra-long longwall face at 1000 m depth (modified from Wang et al.,2019).

Fig.28.Resistance pressure of hydraulic shields for an extra-long longwall face at 1000 m depth (modified from Wang et al.,2019).The numbers with different colors refer to ID of the hydraulic shields at the longwall face.

(2) Pre-tension and its effective transfer into the surrounding rocks are critical for the effectiveness of rock bolt supports.Reasonable pre-tension can effectively restrain the delamination of surrounding rocks,bring the surrounding rock mass into a compressive stress state,and form a pretensioned bearing structure in the anchored zone.

(3) Rock bolting has limited control on the continuous deformation of surrounding rocks in deep roadways,including the elastoplastic deformation and overall extrusion.This overall extrusion is particularly prominent in deep,highly stressed soft rock roadways.The deformation process of each part of the rock bolt also varies,and large local deformation may lead to premature rock bolt failure.Sufficient elongation and impact toughness are critical for effective rock bolting in deep roadways.

(4) Under suitable conditions,rock bolting with high pretension,high strength,high elongation and high impact toughness is an effective measure for coal mine roadways with extensive deformation or rock burst in deep roadways.

(5) For deep roadways where rock bolt and cable support alone cannot effectively control the large deformation of surrounding rocks or burst failure,other support and reinforcement methods or stress relief methods need to be used jointly.

According to the above theory of pre-tensioned rock bolt support in deep roadways,high-strength,high-elongation and high impact toughness rebar materials (Table 5) and large-tonnage and high-elongation cable materials have been developed for deep roadways.The BHTB600,BHTB700 and BHTB800 steel rebars in Table 5 are made of ordinary hot-rolled steel bars,and grain refinement of rebar organization is achieved by medium frequency+ultrasonic induction rapid heating,and continuous heat treatment process of carbon blending+tempering in the martensitic phase transformation zone after quenching to obtain ferrite,martensite and residual austenite complex phase organization.The tensile strength of the rebar with heat treatment is substantially enhanced.Meanwhile,the rebar elongation is not less than 18%.The impact toughness of rebar is also obviously increased with impact absorbing energy amounting to 100-150 J.Largetonnage and high-elongation cable bolts have been developed,with diameters of 22-28.6 mm and breaking forces of 550-900 kN,and their elongation is one times higher than that of traditional cable bolts.

5.2.Grouting reinforcement

For deep roadways with relatively massive rocks,the abovementioned high-strength pre-tensioned rock bolt and cable support can achieve a better support effect.For deep roadways with fractured surrounding rocks,the reinforcement effect of rock bolts and cables might not be guaranteed due to the low bonding strength between the rock bolts and fractured surrounding rocks.Thus,the capacity of rock bolts and cables cannot be fully utilized.Under such conditions,grouting reinforcement is an effective way to control the surrounding rocks in deep roadways.Field experiences have shown that grouting reinforcement of roadways has the following effects:

(1) Grouting improves the strength and stiffness of discontinuities within the surrounding rocks.The strength and deformation of the rock masses are basically controlled by discontinuities.The strength and stiffness of the discontinuities are relatively low,which are prone to opening and sliding,resulting in the reduction of rock mass strength and increase of dilation and increase of the deformation of roadways.Grouting reinforcement can significantly improve the strength and stiffness of the discontinuities and consequently the strength of the surrounding rocks.

(2) Filling and compacting fractures.Under the action of pumping pressure,the grout penetrates and fills some fractures,and closes some fractures that are not filled,which improves the stress distribution state around the fractures and increases the strength of surrounding rocks.

(3) Grouting seals water and isolates air.Grouting in surrounding rocks can effectively seal the channels of flowing water,prevent or reduce the softening of surrounding rocks induced by water,and avoid the significant reduction of surrounding rock mass strength due to the influence of water.At the same time,the fractures are sealed after grouting,which can effectively prevent the surrounding rocks from weathering.

(4) Grouting reinforcement increases the strength of fractured surrounding rocks,and improves the bonding strength of rock bolts and cables,leading to a combined effect of grouting reinforcement and rock bolt reinforcement.

Grouting materials can be generally divided into 3 categories:cement-based,chemical and cement-chemical composite grouting materials.Cement-based grouting materials have the characteristics of high strength,long durability and low cost.But their large granularity makes it difficult to inject the grouts into small fissures of rock masses.Besides,the setting time is not easy to control,and the early strength is low.Chemical materials have characteristics of fast curing and high permeability,but they have disadvantages of high cost,environmental pollution,and potential safety hazards.Inorganic and organic composite grouting materials become the development direction of grouting materials.

Fig.29.Principles and technologies for ground control of coal mine roadways.

Fig.30.Theory and technology system of ground control in deep coal mine roadways.

Fig.31.Reinforcement principle of pre-tensioned rock bolting.

For deep soft rock roadways,under the superposition of high geotress and mining-induced stress,the surrounding rocks contain both large-aperture cracks and small-aperture micro-cracks.These cracks may be isolated and not connected.When grouting is performed,grout can only go into the cracks that are connected to the grouting borehole.Isolated cracks are not connected,leading to poor grouting conditions.To solve this problem,a study on grouting materials was carried out.In order to meet the requirements of“high permeability,high strength,high adhesion and low cost”for the grouting material in the surrounding rocks,an inorganicorganic composite micro-nano grouting material was developed(Guan et al.,2020;Zhang et al.,2020a).A high-pressure fracture grouting process was associated grouting equipment was developed,forming a complete set of high-pressure fracture grouting technology.

The high permeability of the developed grouting material is achieved in the following ways: (1) Reduce the particle size of inorganic grouting material.Use calcium sulfoaluminate minerals with high hydration reactivity as the basic inorganic grouting material,and reduce the particle size through ultra-fine processing,with the material particle sizeD95<7 μm,in order to improve injectability.(2) Improve the wetting of the slurry on the fracture surface.An amphiphilic organic conditioning agent is synthesized,which can significantly reduce the contact angle between water and coal surface and greatly improve the interfacial wettability.(3)Connect isolated cracks through high-pressure fracturing.

The high strength of the grouting material mainly refers to high early strength,which is achieved by three ways:(1)Choose mineral materials with a rapid hydration reaction.(2) Add nano-enhanced materials.Lithium-aluminum hydrotalcite,a nano-enhanced material containing lithium ions,was synthesized to bring into play the synergistic enhancement of lithium ions and nano-materials.(3) Reduce the particle size of the material to increase the reaction area.

High adhesion is achieved in two ways:(1)Improve the wetting of grout on the fracture surface so that the grout is closely bonded to the fracture surface;and(2)Form a chemical bonding interaction between the grout and the fracture surface,which significantly improves the bonding force.The physical and mechanical properties of the inorganic-organic composite micro-nano grouting material developed through the above procedure are listed in Table 6.

Table 5 Mechanical properties of high-strength steel rebars for rock bolts.

Table 6 Properties of micro-nano grouting material (Guan et al.,2020).

A field trial test was carried out at a headgate of Kouzidong coal mine to test grouting reinforcement of the developed inorganicorganic composite micro-nano grouting material.The axial load of the rock bolts in the pillar side wall was recorded before and after the grouting.A significant increase from 37 kN to 145 kN in the axial load was observed.Coal specimens after grouting were extracted and brought to the laboratory for scanning electron microscope (SEM) observation to analyze and evaluate the penetration and bonding of grout in the coal(see Fig.32).The grout could penetrate into the microcracks with an aperture of less than 10 μm.The grout was tightly bonded to crack surfaces (Fig.32b),which greatly increased the reinforcement effect.

5.3.Stress relief method

Stress relief is an efficient way to achieve ground control of roadways subjected to high in situ stresses,high mining-induced stress and rock or coal burst.Commonly used relief methods include drilling,blasting,slotting,and HF in the surrounding rocks.Among these,HF has been a useful measure for stress relief in China’s coal mine practices.The CMB-CCRI has carried out extensive research in the past decades on the development and application of HF techniques for roadway stress relief in regard to the following topics:

(1) HF mechanisms.Theoretical analysis,laboratory tests,and numerical simulations have been performed to understand the initiation and propagation of HFs.The experimental results show that the introduction of a pre-cutting in an HF borehole can only change the HF orientation during the initiation stage.The HF trajectory then returns to the supposed direction perpendicular to the minimum principal stress at a limited distance from the borehole wall.Precutting is thus of minimal engineering significance.Extensive attention has been given to the intersection between HFs and pre-existing fractures.Simulation results indicate that HFs tend to cross pre-existing joints upon high-magnitude fracture cohesion,friction angle,stress difference,approach angle,fluid viscosity,and injection rates (Zhao et al.,2020).Simultaneously propagating fractures can move toward or away from each other,which results in complicated nonplanar geometry and branching of HFs,as shown in Fig.33.An adjacent layer with a low modulus restricts the fracture height growth,whereas the contrast between lateral growth and stress interference will be relatively enhanced (Zhao et al.,2021).These findings further provide insight into the design and optimization of multi-stage HF in horizontal boreholes.

(2) Stress relief mechanism via HF.Although HF has been demonstrated to be an effective way to release high mininginduced stress in field,its mechanism still remains unclear.The HF stress relief mechanism is notably distinct from thatof blasting,which is related to the transfer of stress into deeper rocks because the broken rock cannot sustain stress due to blasting.However,HF does not break rocks like blasting does.Numerical and experimental studies suggest that the two stress relief mechanisms can be applied depending on the location of the HF with respect to the mining-induced stresses.For stress relief at longwall entries,HF is performed in the roof above the entry and pillar and its mechanism for reducing the extensive stress is that the fluid injection generates new HFs and the pre-existing fractures including bedding planes and cross joints are stimulated by increasing the pore pressure and reducing the effective stress acting on the fractures.Shear sliding along HFs and preexisting fractures promotes irreversible and volumetrically distributed changes in the rock mass and consequently relieves stress.If designed appropriately,HF operations in the main roof can substantially reduce the abutment stresses via sliding along the HFs and natural fractures,rather than shifting the abutment pressure to the interior rock mass(Kang et al.,2018b),as shown in Fig.34.Preconditioning is suggested to reduce abutment stresses because HF performed prior to mining has a better performance than postconditioning where HF is performed after mining.HF can reduce the integrity of strong rock layers so that they can collapse promptly as the longwall face advances,which is beneficial for stress relief.The collapse of a localized immediate roof may not substantially affect the front abutment pressure.The continuous collapse of an immediate roof can cause timely fracturing and collapse of the overlying main roof layers and affect the large overburden structure in the longwall face,which thus influences the distribution of mining-induced stress.Continuous HF of the rock layers above a longwall panel can reduce the distance of the initial roof collapse and periodic collapse distance.

(3) HF methods.HF used for underground mining can be divided into two groups: the regional and local HFs.Underground regional HF involves the drilling of long horizontal boreholes in the strong roof above a coal seam prior to mining.The length of the horizontal boreholes is generally at the scale of the mining area,or along the coal mining face to implement fracturing that covers the entire mining area or face.This reduces the mining-induced pressure caused by the hanging of competent roof layers during the mining process.Underground local HF is a method to weaken the hard roof or cut the roof to relieve the pressure in specific locations such as entries and open-off cuts.Each of these two HF methods has its advantages and has been applied in mines under different conditions.

5.4.Strata control using bolting-modification-destressing in synergy

For roadways experiencing severe and/or unfavorable ground conditions(e.g.severe squeezing,highly fractured rock mass,large mining depth,and rock bursts),a single ground support system may not be sufficient to maintain the stability and deformation during excavation.Combined ground control systems are therefore necessary in same cases.The strata control method of boltingmodification-destressing was thus put forward.

Numerical simulations can be used to understand the synergistic effect of the bolting-modification-destressing method for a deep soft rock roadway in Kouzidong coal mine.Fig.35 shows the deformation and damage of surrounding rocks of the roadway under different support patterns.Without support,the roadway was severely damaged and closed due to extreme squeezing.With the rock bolt and cable support,the deformations of roadway walls and roof were significantly reduced.The high mining-induced pressure on the roadway was released mainly through the deformation of the floor,causing extensive floor heave.The use of rock bolt and cable support failed to control the severe deformation of the surrounding rocks.By grouting the roadway surrounding rocks,the bearing capacity of surrounding rocks was significantly improved.However,due to the high rock stress and the extensive mining disturbance,the wall-to-wall convergence still reached 1.2 m.With destressing by HF performed within the roof,the high abutment pressures,roof sag,and wall-to-wall convergence were effectively reduced.

Fig.32.SEM images of coal specimen after grouting.

Fig.33.Simultaneous growth of multiple HFs under different stress regimes.The HF aperture was truncated at 0.001 mm and 0.5 mm (modified from Zhao et al.,2021).

Fig.34.Effect of HF on the release of mining-induced pressure: (a) No HF treatment;and (b) With HF treatment (modified from Kang et al.,2018b).

Based on the above analysis,the synergistic control principle of the bolting-modification-destressing method mainly involves three aspects(Kang et al.,2020).First,high pre-tensioned rock bolts and cables can reduce the shallow deviatoric stress and stress gradient of the surrounding rocks,control the discontinuous and uncoordinated dilatant deformation of the surrounding rock mass in the bolted zone,and increase the surrounding rock strength.Second,modification by fracture grouting can improve the strength and integrity of the two side walls,and increase the bonding strength of wall bolts and cables.This not only reduces the deformation of the side walls,but also improves the ability of the roadway walls to stabilize the roof.Third,HF within the roof can reduce the length of the cantilever.In addition,sliding along the hydraulic and preexisting fractures can greatly reduce the mining-induced stress applied on the roadway.Through the combination of high-strength supporting of pre-tensioned rock bolts and cables,rock mass modification with fracture grouting,and distressing through HF,the deformation and failure of 1000-m deep roadway with soft rock can be successfully controlled.

5.5.Cases of roadway ground control in deep coal mines

5.5.1.Rock bolting and grouting technology for main roadways of over 1000-m depth

The Xinwen coal area is a typical deep mining area in China,with a maximum mining depth of 1510 m.The depth of the-1180 main return roadway at Huafeng mine was more than 1200 m.High geostress led to severe damage of the roadway as shown in Fig.36.It is rather difficult to control the extreme deformation using conventional support methods.Based on the results of field measurements and numerical simulation,a combined support system of high pre-tensioned high-strength rock bolts and cables,grouting and shotcrete was proposed to support the-1180 main ventilation roadway,as shown in Fig.37 (Kang et al.,2015b).

The resin bolts for roof and side were 22 mm in diameter and 2.4 m in length,with a capacity of 300 kN.The row spacing of the rock bolts was 0.9 m,with 15 rock bolts in each row.Cable installation and grouting were carried out in the same borehole.The cables had a diameter of 22 mm,with a capacity of 560 kN and a length of 4.3 m.The head of the cables was anchored with resin,and the reminding section was anchored with cement grouting.The row spacing of the cables was 0.9 m with 5 cables in each row.The grouting material was #525 ordinary Portland silicate cement with XPM nano-grouting additive.The roadway floor was supported by cables and grouting.The floor cables were the same as the roof cables.A“birdcage”was installed in the anchorage section to improve the bonding strength of the floor cables anchored with resin and grouting as the roof cables.

After adopting the above support technology,the displacement of the surrounding rocks in the-1180 main return roadway did not exceed 100 mm,and the shallow(2.3 m)and deep(4 m)departures of the roof were 4 mm and 7 mm,respectively.It showed that this support method could effectively control the deformation of deep roadway with the mining depth greater than 1000 m and maintain its long-term stability.

5.5.2.Ground support technology for deep entry retained along gob

The Huainan mining district is a typical deep coal mining district with soft rock masses.A field trial was conducted at the tailgate of the 5121B10 longwall face of Xieyi coal mine.The geological and engineering conditions of the longwall face and the deformation and damage of the tailgate with the original support system are shown in Section 4.1.3.According to the deformation and damage characteristics of the surrounding rocks in the deep soft rock tailgate,a support technology was proposed,including the entry support and side wall support(Kang et al.,2010a).The trapezoidal cross-section of the entry was selected according to the geological conditions.The coal measure strata were inclined with an angle of around 22°.To maintain the integrity of the roof for a favorable ground control condition,the entry was excavated along the roof.The entry support consisted of basic and enhanced supports.The basic support involved high pre-tensioned,high-strength and highelongation rock bolts and cables,as shown in Fig.38.The resin rock bolts were 22 mm in diameter and 2.4 m in length,with a spacing of 0.9 m × 1 m.The row spacing of the side bolts was 1 m.The roof cables were 22 mm in diameter and 6.3 m in length,and two cables were in a row at a spacing of 2 m.The enhanced supports were individual hydraulic props with hinged top beams,installed from 60 m in front of the working face to 100 m behind the face.Paste filling was adopted for the side support.The materials of the paste filling were composed of cement,pulverized fuel ash,pebble,sand and admixture with 5-d compressive strength of 10 MPa and 28-d strength of 14 MPa.The width of the filled side wall was 2.5 m and the height was 1.6 m.

Fig.35.State of roadway displacement contained by various ground control patterns: (a) No support;(b) Rock bolts and cables;and (c) Bolting+modification+destressing.

Fig.36.Deformation and damage of a main roadway with depth over 1200 m in Huafeng coal mine: (a) Severe roof sag and wall-to-wall convergence;and (b) Severe floor heave (U-shaped steel arches inserted into the floor).

Fig.37.Rock bolting pattern for the-1180 main return roadway in Huafeng coal mine(unit: mm).

Utilizing the above-mentioned support method,the deformation of deep entry retained along gob in Xieyi coal mine was effectively controlled during the service period (Fig.39),and the surrounding rocks and the filled side wall were stable,which met the requirements of safe mining production.

5.5.3.Ground support technology for deep burst-prone roadways

Yima mining district is a typical burst-prone coal mining district in China.Coal bursts greatly threaten safe coal production.The 21220 longwall face in the Changcun coal mine has a depth more than 800 m.Coal bursts occurred frequently in this mine,as described in Section 4.1.4,leading to severe damage to roadways(Fig.23).To solve this problem,a supporting principle for the burstprone roadway was proposed based on the roadway deformation in addition to damage and support mechanism of burst-prone roadways.In principle,rock bolts and cables are the fundamental support elements,and can be used in combination with stand support elements such as steel sets and hydraulic piles.Destressing can also be considered.High pre-tensioned,high-strength,high-elongation,high impact toughness and fully encapsulated rock bolts are considered to be suitable for burst-prone ground (Kang et al.,2015a).

According to the aforementioned principle,a support system for the 21220 headgate was proposed,as shown in Fig.40.The rock bolts had a diameter of 22 mm,a length of 2.4 m,a tensile capacity of 300 kN,an elongation of 22%,and an impact absorbing energy of 120 J.The rock bolts were fully encapsulated and pre-tensioned.The cables were 22 mm in diameter,4-6 m in length,1.8 m spacing,and 0.9 m row spacing.The 36U steel sets were tri-centric arches with 1.2 m spacing.Hydraulic sheds with a working resistance of 2200 kN were installed.Drill-and-blast method was used at the corners of the two walls for distressing purpose.

Fig.38.Bolting pattern of the tailgate in Xieyi coal mine (unit: mm).

Field monitoring data showed that the headgate deformation was dominated by the wall-to-wall convergence and floor heave.The surrounding rock-support system was damaged to some extent due to burst events.The maximum roof sag,wall-to-wall convergence and maximum floor heave were 110 mm,550 mm and 700 mm,respectively.The support condition of the headgate was significantly improved and the deformation of the surrounding rocks was effectively controlled.Rehabilitation was not needed during the service period of the headgate,except that dinting was performed.The maximum energy of a single event was less than 107J.Premature failures of rock bolts and cables were not observed.

5.5.4.Bolting-modification-distressing ground control technology for 1000-m deep longwall entries with soft rocks

Fig.39.High-strength bolting states in Xieyi coal mine.

Longwall entries at Kouzidong coal mine are typical deep roadways with soft rock mass,high mining disturbance and large deformation.The deformation and failure pattern of the entries are discussed in Section 4.1.2.In order to solve this kind of extremely difficult problems in roadway support,a strategy was proposed that combines rock bolting,rock mass modification,and destressing(Kang et al.,2020,2021a),as shown in Fig.41.Rock bolts and cables were the fundamental support elements (Fig.41a).The rock bolts were BHTB700 super-high-strength rebar with 22 mm in diameter and 2.4 m in length,with spacing of 0.9 m×0.9 m.The cables were 22 mm in diameter.The roof cables were 7.2 m long,with spacing of 1.6 m × 0.9 m.The wall cables were 4.2 m long,with spacing of 1.35 m × 0.9 m.Splitting grouting with high pressure was carried out in fractured coal pillars before the heading face (Zhang et al.,2020b).Two grouting boreholes were arranged in a daily cycle(Fig.41b).The grouting material was the modified micro-nano inorganic-organic composite material as mentioned above.The water-cement ratio was 0.8-1.When the grouting pressure amounted to 9 MPa,fracture grouting initiated.HF was performed in the hard fine sandstone in the range of 37.1-45.3 m deep into the roof.The layout of the boreholes for hydraulic fracturing is shown in Fig.41c.Two groups of boreholes with a spacing of 10 m and a depth of 67 m were arranged on both sides of the entry.The boreholes were sealed section by section,starting from the bottom of the borehole,until the hard roof was completely fractured.

Field monitoring data suggested that in contrast to the initial support,the entry convergence was reduced by more than 50%after the bolting-modification-destressing pattern was applied,and the rate of premature failure of rock bolts and cables was reduced by 90%.The combined ground control system significantly mitigated the extreme squeezing conditions,as shown in Fig.42.

6.Ground control at deep longwall faces

Hydraulic support is a key support measure for controlling the mining pressures at a longwall face using the fully mechanized mining method.The design of hydraulic support type and parameters is very important for successful ground control in deep working faces (Wang et al.,2021b).

Fig.40.Support pattern for the 21220 headgate in Changcun coal mine (unit: mm).

According to hydraulic support and fracturing characteristics of overlying strata on the working face,a dynamic model that can consider strength,stiffness and stability of the hydraulic supports and surrounding rocks under dynamic pressure of deep mine was established(Fig.43).A method for analyzing the strength,stiffness and stability of supports suitable for deep longwall faces was also proposed (Wang and Pang,2015).

6.1.Principle of interaction between hydraulic supports and surrounding rock masses

A method was proposed to determine the six controllable parameters of hydraulic supports for effective control of surrounding rocks based on the coupling support principle of hydraulic supports and surrounding rocks.The six parameters are initial support pressure,working pressure,location of combined forces,horizontal force,length of support beam,and moving speed.The formula for calculating the working resistance pressure of traditional hydraulic support is modified as (Xu et al.,2015):

wherePis the working resistance pressure(MPa),?is a correction coefficient,γ is the roof unit weight(kN/m3),Mis the mining height(m),andKpis the crushing and expansion factor of the roof.

6.2.Design and development of high-reliability support system for deep super-long working face

Fig.41.Pattern of the combined ground support-rock mass modification-destressing strategy:(a)Layout of rock bolts and cables;(b)Layout of grouting boreholes;and(c)Layout of HF boreholes (modified from Kang et al.,2020).

Fig.42.State of entry controlled by rock bolting-grouting-destressing.

Fig.43.Coupling support principle of hydraulic shield and surrounding rock mass.

The hydraulic supports in underground coal mines in China can be classified as chock,shield and chock-shield types according to their structures.The geometric and mechanical parameters of hydraulic supports include height,width,setting load,working load,support intensity,floor specific pressure,pump pressure.The design of hydraulic supports should be conducted as per sitespecific conditions.The “three-coupling” parameter refinement design method of the support system was proposed.The design and development of the four-column non-isointensity hydraulic support with length of 7 m were realized.The support can be used for the ultra-long working face at 1000 m depth with large mining height (Fig.44).The hydraulic support has a working resistance of 18,000 kN and an overall support strength of 1.73-1.78 MPa.The diameter of the front column of the hydraulic support is 440 mm,and the diameter of the rear column is 360 mm.The non-isointensity hydraulic supports can solve the difficulties upon underframe sinking and frame moving.The support can partially protect the working face with large mining height,in addition to broken roof and leaks of top coal.A highadaptability,high-reliability and high-stability support system was then formed for the working face with complex ground conditions at 1000 m depth.

An intelligent mining technology of “perception-predictiondecision”under complex ground conditions,such as 1000-m deep mines with ultra-long working face,was proposed to realize environmentally adaptive technology.The technology can effectively improve the surrounding rock mass stability and mining efficiency(see Fig.45).

6.3.Ground control technology for strong roofs

6.3.1.Deep borehole blasting technology

The impact loading generated by blasting is applied to the rocks around borehole wall.The rock is subjected to dynamic compressive stresses.The expanding cavity,crushing zone,fractured zone,and vibration zone are produced from the inside outwards,thus destroying the integrity of surrounding rocks.This is suitable for strong roof conditions where fissures are not well developed.When the coal seam is gas-rich or has the tendency to explode,it is necessary to take measures to prevent coal dust and gas explosion when implementing an explosive blast.Fig.46 shows the layout of blasting holes for hard roofs at the longwall face of the No.12 coal mine,a 1000-m deep mine in Pingdingshan mining district (Zhao et al.,2015).Based on the characteristics of thick and hard rock deposit on the longwall face,blasting was implemented along the inclined direction of the working face to promote the breakage and collapse of the roof behind the longwall face.Field observation suggested that the blasting reduced the first roof weighting distance and periodic roof weighting intervals,and mitigated the risk of dynamic disasters induced by sudden caving of a large area of the overhanging roof behind the longwall face.

6.3.2.HF treatment

In addition to reduce mining-induced stresses,HF can also be used to precondition strong roofs by injecting high-pressure water into hard rock to create fractures or reopen natural fractures,and to weaken the strong roof to form a fractured network that can collapse regularly.Numerical result using a hybrid finite element/discrete element method showed that after HF treatment,the strong roof above the gob can cave and collapse more easily,leading to a smaller periodic roof weighting interval,as shown in Fig.47.

HF as a means to precondition a strong roof can be performed on the ground surface or underground.The surface HF method is similar to that in the petroleum industry,in which vertical or horizontal wells are drilled from the surface into the target layer to fracture the hard rock formation in sections (Yu et al.,2019).Underground HF is similar to the aforementioned approach for stress relief of a roadway via HF and is divided into regional and local fracturing.The former is used to deal with a large area of hard roof above the gob,while the latter is used for local sections such as the initial caving of the main roof.

Fig.44.Photo of 7-m non-isointensity four-column hydraulic shield with large mining height.

Fig.45.Ultra-long longwall face at Kouzidong coal mine.

Fig.46.Layout of deep-hole pre-splitting blasting for 1000-m deep longwall face with a thick and hard roof (Zhao et al.,2015).

6.4.Ground control technology for fractured coal and roofs

Grouting can effectively reinforce fractured rocks surrounding roadways as discussed previously.Grouting is also an effective way to control the fractured roof and coal seam of a longwall face.Deep longwall faces with large mining heights are more likely to experience coal face bulking and even roof collapse.Chemical grouting can provide fast and effective reinforcement owing to the strong permeability,fast curing,and high strength(Wu et al.,2012),and is are widely used for ground reinforcement of longwall faces.

Deep-hole pregrouting of a longwall face from entries is another promising method.The drilling location and parameters are determined according to the geological conditions of the longwall face,and the grouting reinforcement is implemented before the longwall face crosses the fractured zones in the coal seam.Deephole pregrouting does not affect the normal production of the longwall face and cement-based grouting materials can be used.This significantly reduces the reinforcement cost(Wang,2018)and good technical results and economic benefits have been achieved.

7.Conclusion and future work

Fig.47.Numerical prediction of HF treatment on the preconditioning of a strong roof above a mined-out area: (a) Without HF treatment;and (b) With HF treatment.

Significant breakthroughs in ground control technology have been made for deep underground coal mining practice in China in recent decades.This has improved our understanding of the geological and geomechanical characteristics of coal measures,as well as evolution of mining-induced stress and structures around deep roadways and longwall faces.The concept of ground control in roadways has progressed from passive support to active control of surrounding rocks and support elements.The concept of ground control using pre-tensioned rock bolts,rock mass modification,and destressing is widely recognized.Ground control design has evolved from static and empirical designs to dynamic,informational,systematic,and scientific designs.Support materials continue to develop towards high-strength,high-stiffness,highelongation and high impact toughness.However,the geological and mining conditions of China’s coal seams are complex.Large differences exist among coal mining districts in terms of mining concepts,technology,equipment,and management.There is still a large space for improvement,especially for deep coal mines with unfavorable and complicated ground conditions.

(1) The capacities for in situ testing and analysis of the geomechanical characteristics of coal seams should be strengthened.A new classification method should be developed for the rocks surrounding coal mine roadways.This method can involve quick and accurate in situ measurements of geological and geotechnical parameters of the roadway with advanced techniques.The measured data can then be evaluated using machine learning method to output a classification of the geological condition and suggestions for ground support pattern and parameters.

(2) Although a large number of monitoring,theoretical analysis,and numerical studies have been conducted to evaluate mining-induced stresses around deep longwall faces and roadways,associated technology and instruments that can monitor mining-induced stress in a long-term,stable,and effective manner are still lack.The impact of the rock structures that form in the overlying strata above the gob when using different coal mining methods and parameters on the stability of the roadway surrounding rock mass is also not fully understood.Systematic research on mining dynamics is required,including rock stress measurement,monitoring,simulations,and theoretical analyses.

(3) Effort should be devoted to studying the mechanisms of mining-induced fracturing and deformation in long and deep longwall faces.Extensive attention should be paid to the precursors of typical types of roof-involved hazards to achieve a real-time warning system for intelligent decisionmaking and control of roof-involved hazards.

(4) The interactions between pre-mining stress,mining-induced stress,and support stress should be investigated,in addition to the theories and technologies for ground control of roadways under extremely difficult conditions,such as great depth (over 1000 m),burst-prone tendency,strong mining disturbance,and extremely soft and fractured surrounding rocks.

(5) With rapid development of digitalization and intelligent construction of China’s coal mines,an intelligent monitoring and early-warning system of mining pressure in longwall faces and roadways should be developed.The big data system of deep mining and strata control should be established,for the purposes of accurate prediction,prevention and control of disasters in deep mining.

It should be noted that the review is based only on Chinese underground coal mining practice.The developed materials,technologies,equipment and proposed theories,however,are of valuable reference to deep coal mines around the world.

Declaration of competing interest

The authors declare that they have no known competing financial interests or personal relationships that could have appeared to influence the work reported in this paper.

Acknowledgments

This work has been supported by the National Key Research and Development Program (Grant No.2017YFC0603000),which was jointly completed by the Coal Mining Research Branch of CCRI,China University of Mining and Technology (Xuzhou and Beijing),Henan Polytechnic University,and Xinji Energy Company Limited of China Coal Energy Group.This work was also supported by the National Natural Science Foundation of China (Grant No.51927807).The authors would like to thank all persons for supporting this work.