Research ArticleStrong Ground Pressure Mechanism and Control at the LongwallTop Coal Caving with a Single Key Stratum in Goaf

Ke Yang12345 Xiaolou Chi 1345 Wenjie Liu1345 Litong Dou1345 and Zhen Wei1345

1State Key Laboratory of Mining Response and Disaster Prevention and Control in Deep Coal MinesAnhui University of Science and Technology Huainan Anhui 232001 China2Institute of Energy Hefei Comprehensive National Science Center Hefei Anhui 230031 China3Key Laboratory of Mining Coal Safety and Efficiently Constructed by Anhui Province and Ministry of EducationAnhui University of Science and Technology Huainan Anhui 232001 China4National amp Local Joint Engineering Research Center of Precision Coal Mining Anhui University of Science and TechnologyHuainan Anhui 232001 China5School of Energy and Safety Anhui University of Science and Technology Huainan Anhui 232001 China

Correspondence should be addressed to Xiaolou Chi chixiaolou163com

Received 30 May 2020 Revised 20 July 2020 Accepted 4 August 2020 Published 24 August 2020

Academic Editor Xianjie Hao

Copyright copy 2020Ke Yang et al+is is an open access article distributed under the Creative CommonsAttribution License whichpermits unrestricted use distribution and reproduction in any medium provided the original work is properly cited

A strong ground pressure in the multiseam environment manifested by rib spalling and roadway deformation at the fullymechanized working face was assessed by a comprehensive combination of field measurements physical simulations andtheoretical analysis for two coal seams in the Buertai Coal Mine in China A structural model of overlying stratum collapse at theworking face with the key stratum breaking instability was proposed the mechanism of strong ground pressure at the longwall topcoal caving working face with a single key stratum in goaf was identified and respective control countermeasures were developed+e latter implied the directional hydraulic fracturing for supporting the key stratum-surrounding rocks which effectivelyreduced the cyclic weighting intensity and weighting interval in the working face with a single key stratum in the goaf+eworkingface cyclic breaking interval was assessed at 30m After the key stratum collapse soft rocks underwent synergistic deformation anda cutting-type failure +e goaf effect on the hydraulic support resistance in the fully mechanized working face was assessed andcutting blocks from the overlying stratum collapse were identified as the main sources of strong ground pressure

1 Introduction

Coal consumption in China accounted for 590 of the totalenergy consumption in 2018 which makes the issues of safeand cost-efficient coal mining quite topical One of the coalmining hazards is that under coal-forming conditions hardstrata overlying the working face increase their groundpressure and are prone to collapse Moreover as the miningheight increases the overlying strata undergo extensivemigration and hard strata exert even a more conspicuousand complex impact on the working face [1 2] In particularit was revealed that hard rocks in the overlying strata usuallycontrol the breaking and migration of the local strata whilethe size and breaking mode of broken blocks have a decisive

impact on the intensity of ground pressure manifestationMany scholars have studied the breaking-related disaster-triggering pattern in light of the breaking features andmovement mode of the key strata (hard rocks) +e effects ofcutting structures from the key cantilever beam and theO-type breaking failure of the key strata on the groundpressure manifestation at the working face were analyzed[3ndash7] +e above studies formed the basis of the stratacontrol theory However due to the coal seam enrichmentconditions multiseam coal beds are usually present in thestrata and affect their load distribution +erefore loadtransfer mechanisms under different coal pillar widthsduring multiseam mining were also extensively investigated+e relationship between the height of the transmissive

HindawiShock and VibrationVolume 2020 Article ID 8835101 12 pageshttpsdoiorg10115520208835101

fracture belt and coal seam spacing was established and thedynamic disaster occurrence mechanism under the influ-ence of mining-induced stress at the working face wasidentified [8ndash14] +e design of rock support system underrockburst condition was proposed by Kaiser and Cai [15]However quite a few studies were focused on the key stratabetween the seams in multiseam mining or on the collapsefeatures of the overlying strata with synergistic breaking+is paper analyzes the problem of strong ground pressureat the longwall top coal caving (LTCC) working face in thegoaf using a comprehensive combination of field mea-surements physical simulations and theoretical analysis+e collapse of the overlying stratum during the extractionof the working face is studied in detail +e mechanism ofstrong ground pressure manifestation at the working facewith a single key stratum between the seams in the lowergoaf area is identified Based on the directional hydraulicfracturing technology [16ndash21] the fracture scheme for hardrocks is proposed which has lucrative technical and eco-nomic prospects +e results obtained provide theoreticaland technical guidance for safe extraction of coal minesunder similar engineering and geological conditions

2 Project Overview

21 Geological Conditions and Mining Layout +e BuertaiCoal Mine is located in the southeast of Ordos City InnerMongolia Autonomous Region of China At the workingface 42017 of the Buertai Coal Mine the coal seam 4-2 is themain extractable seam with a burial depth of 339ndash460mthickness of 65m (090ndash768m) the dip angle of 1ndash3deg andworking face width of 300m +e lithologic features of thecoal seam roof and floor are listed in Table 1 +e coal seam2-2 overlying the coal seam 4-2 with a distance of 65m hasa thickness of 082ndash580m and burial depth of 212ndash360mBoth seams have simple structures and well field profiles aswell as stable horizons +e longwall top coal caving (LTCC)method is applied to the working face with a mining heightof 35m and a caving height of 3m +us the mining-to-caving ratio is 1 086 and the average daily advance speed is10md Based on the drill column at the working face 42107and the key stratum theory [22 23] the 22m-thick siltstonebetween the two seams is identified as the key stratum +egeographic location of the working face and the mininglayout are shown in Figure 1

22 Ground Pressure Behavior Figure 2 shows the fieldmeasurement data on ground pressure distribution+e firstweighting interval of the working face 42107 is about 100mand the first cyclic weighting interval is about 30mZFY210002539D two-column cover-type caving hydraulicsupports are used for the working face 42107 +e openingpressure of the safety valve of the supports is about 46MPaWhen the working face 42017 is mined at 130m from thesetup entry the support working resistance at the workingface rises to 50ndash52MPa as shown in Figure 2 At thismoment the safety valve of the supports opens and there is asevere rib spalling at the working face+e floor vibrates and

the shearer bounces Such strong ground pressure mani-festations as roadway floor heave and deformation in thesegments with advanced support occur (Figure 3) whichjeopardize the coal mining operation and safety To ensurethe subsequent extraction safety of the working face oneneeds to clarify the occurrence mechanism of strong groundpressure and to develop effective control countermeasuresagainst its hazardous manifestations

3 Mechanical and Similar Physical Models ofthe Overlying Stratum Breaking

31 Mechanical Model of the Key Stratum Breaking IntervalAs shown in Figure 1 there is a 22m-thick key stratumbetween coal seams 4-2 and 2-2 According to the keystratum theory the key stratum physical-mechanicalproperties control the cyclic breaking interval After thetensile stress in a cross section exceeds its tensile strengthvalue microcracks start to propagate within the key stratumand their coalescence into macrocracks finally results in thekey stratum breaking and collapse+e stress state before thebreaking and rotary instability of the key stratum can bereduced to a pure bending of a cantilever beam as shown inFigure 4 where M is the bending moment kNm q is thedistributed compressive stress MPa h and b are the rect-angular key stratum thickness and width respectively mymax is the maximum roof suspension distance m

+e relationship between the tensile strength andbreaking interval of the key stratum can be derived as fol-lows +e maximum tensile stress σmax and maximumbending moment M are linked as

σmax Mh

2lz (1)

where h is the beam thickness and lz is the moment of inertiaof the rectangular cross section with respect to axis Z m4+is implies that the maximum normal stress in a rectan-gular cantilever beam in pure bending occurs at the furthestposition from the neutral axis Values ofM and lz are derivedas follows using the same definitions as in Figure 4

M 1113946ymax

0qy dy (2)

lz bh

3

12 (3)

By substituting formulas (2) and (3) into (1) the rela-tionship between the cyclic breaking interval of the keystratum and its tensile strength is obtained

ymax h

σt

3q

1113970

(4)

where σt is the tensile strength of the key stratum MPa

32 Breaking Interval of the Key Stratum According toformula (4) lower key stratum tensile strength values cor-respond to smaller values of the maximum suspension roof

2 Shock and Vibration

distance During the advancing of the working face a certainroof suspension occurred due to excessively large tensilestrength of the key stratum When the ultimate caving

interval was reached in the key stratum the latter broke andthere was the synergistic breaking of the soft rocks above itAfter breaking the rock strata load was transferred to

Table 1 Lithologic features

Rock formation Lithology +ickness (m) Lithologic characteristics

Overburden

Siltstone 80 Gray silty structure semihard homogeneous bedding and quartz-based compositionSandy mudstone 120 Gray argillaceous structure and semihard

2-2 coal 35 Black blocky structure weak bituminous sheen and small amount of pyriteSandy mudstone 190 Gray argillaceous structure semihard and coal-bearing wire

Key stratum Siltstone 220 Gray silty structure hard homogeneous bedding and quartz-based compositionOverburden Sandy mudstone 240 Gray argillaceous structure and semihardCoal seam 4-2 coal 65 Black blocky structure and weak bituminous sheenFloor Sandy mudstone 50 Gray argillaceous structure semihard and coal-bearing wire

Study area (buertai mine)

Overburden caving

Key stratum2-2 coal seam goaf

4-2 coal seam

Inner mongolia autonomous region

China

Ordos

Figure 1 Mine location and working face layout

20

40

60

Wor

king

resis

tanc

e (M

Pa)

Advance distance (m)

100

120130

80

40

20

60

01

25Support number50

75100

125 150

Ground pressure behavior 64MPa

60MPa

56MPa

52MPa

48MPa

44MPa

40MPa

36MPa

32MPa

28MPa

24MPa

20MPa

16MPa

Figure 2 Morphology of borehole walls at different heights

Shock and Vibration 3

supports and coal bodies +e larger the caving interval ofthe key stratum the larger the load acting on the supports

According to the engineering and geological conditionsof the working face 42107 the key stratum thickness was22m its tensile strength was 54MPa and the load exertedby the overlying stratum on the key stratum was 10MPaUsing formula (4) the cyclic breaking interval of the keystratum was assessed as 295m

4 Similar Physical Simulation Test

41 5e Similar Physical Model +e similar physical modeladopted the burial depth thickness and mining-to-cavingratio of the coal seam 4-2 of 460m 65m and 1 086respectively In situ coal samples were collected by coredrilling and used to fabricate cylindrical specimens formechanical tests Considering the discontinuity of rockmaterials the joint length and continuity factor were set to07 +e mechanical parameters of the coal seam 4-2 roofand floor rocks are listed in Table 2

+e geometric similarity ratioCD between the simulationand field-measured parameters was set to 1 100 +e sim-ilarity ratios of the volumetric weights of rocks and coal wereCc-Rock 1 156 and Cc-Coal 1 1 respectively Based on alarge number of tests and simulations the ground pressuresimilarity theory was developed [24 25] which implies thefollowing similarity ratios of model stresses elastic moduliloads and time

According to the similarity theory the following rela-tionship exists between the physical model and geologicalprototype

Cσ CcCD (5)

where σ is stress and Cσ is the stress similarity ratio c is thebulk density Cc is the bulk density similarity ratio D isdefined as length and CD is the geometric similarity ratio

In addition

CD CδCε (6)

where δ is the displacement Cδ is the displacement similarityratio ε is the strain and Cε is the strain similarity ratio

Finally

CE CσCε (7)

where E is the elastic modulus and CE is the elastic modulussimilarity ratio

Similar physical simulations also use such parameters astensile and compressive strength values (σt and σc) cohesiveforce (c) internal friction angle (β) Poissonrsquos ratio (v) andfriction coefficient (μ)

Insofar as parameters of the physical model are equal tothose of the geological prototype the similarity ratios ofstrain internal friction angle friction coefficient andPoissonrsquos ratio are equal to unity

Cε 1

Cβ 1

Cμ 1

C] 1

(8)

Formulas (5)ndash(8) yield the following equality

Cσ CE Cc Cσc Cσt

(9)

For the above values of Cc-Rock 1 156 and Cc-Coal 1 1we get the following stress and elastic modulus similarityratios

Cσ CE CcCD 156 (10)

Rib heave

Floor heave

(a)

Rib spalling

(b)

Figure 3 Strong ground pressure manifestations

b

h

q

M

XY

Z

ymax

Figure 4 Mechanical model of a cantilever beam pure bending

4 Shock and Vibration

+en the load similarity ratio is

CF CcCD3 156 times 106 (11)

+e time similarity ratio is

Ct CD

1113968 10 (12)

Using the obtained relation between stress and elasticmodulus CσE 1 156 the physical-mechanical parametersof the stratum under similar simulation were determinedand the initial mix ratios of similar materials were deter-mined from the similar theory [26 27] Materials used forsimilar simulations included fine sand with the particlediameter range of 025ndash035mm lime and gypsum In orderto obtain the appropriate elastic modulus of the materialfor similar simulation a large number of mechanical testswere performed for the collected specimens +e mixratios of rock similar materials are listed in the lastcolumn of Table 1 First the initial mix ratios of similarmaterials were preset +en weighted mixing was per-formed for similar materials (fine sand lime and gyp-sum) which were used to prepare standard columnarspecimens After curing for seven days these specimenswere subjected to mechanical tests +e final mix ratios ofsimilar materials were determined after screening andcomparative analysis

+e specification of the similar simulation device was asfollows length times width times height 3000 mm times 400 mm times

2000mm Fine sand was used as aggregate lime and gypsumwere used as a cementing agent and water was used as abonding agent Based on the enrichment status of theworking face 42107 in the Buertai Coal Mine the modelspecificationwas as follows lengthtimes widthtimes height 3000mmtimes 400mmtimes 1500mm +e laying thickness of the coal seam2-2 was 35mm and that of the overlying stratum of thecoal seam 2-2 was 700 mm +e laying thickness of thecoal seam 4-2 was 65m and that of the stratum betweencoal seams 2-2 and 4-2 was 650mm the layingthickness of the floor of the coal seam 4-2 was 50mm In

the simulation experiment the advancing distance alongthe orientation of the working face was 2200 mm thedistances from the model left and right boundaries to thesetup entry and coal seam stop mining line were bothequal to 400 mm To achieve uniform loading a multi-function adjustable physical simulation loading devicewas installed above the testing device which partiallycompensated the gravity-induced load of the overlyingstratum by the imposed load 005MPa +ere are 5 mea-suring lines of displacement arranged in the physical model+e physical model is depicted in Figure 5

+e time similarity ratio was the ratio of time periodsrequired for excavations in the physical model and the actualoperation at the working face Since the roof movement isconsiderably affected by the extraction speed of the workingface the ground pressure manifestation pattern shouldchange dramatically with the extraction speed variation atthe working +erefore the time similarity ratio was used todetermine the time needed for excavation in the simulationmodel Based on the time and geometric similarity ratioseach day of excavation at the real working face correspondsto 24 h of excavation in the physical model+e 38 workingschedule is adopted in the Buertai Coal Mine with two shiftsper day To be specific every shift is responsible for exca-vation for 50mm in the physical model (excavation for 5min the actual working face) with two shifts per day +at isthe working face in the model advanced by 100m per 24 hwhile that at the actual working face advanced by 10m per24 h

42 Breaking Characteristics of Overburden After the ex-cavation of the coal seam 2-2 was completed its overlyingstratum broke and vertical cracks formed on both sides ofthe goaf +ere was general bending subsidence of theoverlying stratum Due to the mining-induced influence ofthe coal seam 2-2 the overall strength of the overlyingstratum decreased and the physical-mechanical propertiesof rocks became close to those of soft rocks as shown in

Table 2 Physical-mechanical parameters of coal rocks and mix ratios of rock similar materials

NameStrata

thickness(m)

Tensilestrength(MPa)

Elasticmodulus(GPa)

Cohesion(MPa)

Compressive strength Bulk density Mix ratio

Prototype(MPa)

Model(MPa)

Prototype(kgm3)

Model(kgm3)

Sand gypsum lime

Sandymudstone 500 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 80 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 120 157 634 36 3263 021 2100 1350 7 7 3

2-2 coal 35 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 190 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 220 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 240 157 634 36 3263 021 2100 1350 7 7 3

4-2 coal 65 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 50 157 634 36 3263 021 2100 1350 7 7 3

Shock and Vibration 5

Figure 6(a) After the breaking of the key stratum betweencoal seams 2-2 and 4-2 there would be a synergisticcollapse of the overlying stratum of coal seam 2-2(equivalent to soft rocks) and key stratum When theworking face 42107 advanced to 85m the lower immediateroof began to fall along with the excavation while theimmediate upper roof and key stratum remained intact Atthis moment the breaking angle of the lower immediate roofwas about 52deg as shown in Figure 6(b)

As the working face continued to advance the de-formation of the immediate upper roof and key stratumgradually increased and cracks began to develop Whenthe working face advanced to 100m the 22m-thick keystratum broke in the rear upper part of the coal wall alongthe working face and the breaking angle was about 60deg asshown in Figure 6(c) At this time the working face wassubject to the first weighting which was consistent with anincrease in the support load when the working face ad-vanced to 100m in Figure 2

+e bending deformation of the key stratum and bedseparation space below it began to increase with the workingface advance First transverse cracks appeared in the keystratum +en vertical cracks appeared in the middle of thebed separation space exhibiting a bottom-up developmentWhen the working face advanced to 130m vertical crackspenetrated the entire key stratum +e key stratum brokeand as a result the immediate roof key stratum and load-bearing layer underwent the first cutting as a whole in therear upper part of the coal wall along the working face +ecutting angle was about 64deg as shown in Figure 6(d)Meanwhile there was a synergistic movement of the keystratum and the overlying stratum of the coal seam 2-2where the extraction was completed (equivalent to softrocks)

As the working face continued to advance after theextraction of coal seam 2-2 was completed (equivalent tosoft rocks) the key load-bearing and overlying strataexhibited a cyclic synergistic cutting-type breaking +ecutting position was found in the rear upper part of the coal

wall along the working face +e cyclic cutting angle wasclose to the initial cutting angle as shown in Figures 6(e) and6(f ) +e overlying stratum underwent a cyclic cutting-typebreaking as a whole Two adjacent cutting blocks were inclose contact with each other squeezing each other andforming a hinged structure+is structure subsided slowly asthe overlying stratum broke until it came into full contactwith gangues in the goaf

43 Displacement Characteristics of Overburden Figure 7shows the overburden displacement curve in the miningprocess of number 4-2 coal seam It can be seen fromFigure 7 that the overlying strata of number 4-2 coal seamexperienced a continuous dynamic subsidence andmovement process during the mining process +e sub-sidence trend of the overlying strata is nonlinear and themovement form is asymmetric Specifically when theworking face is advanced to 85m the immediate roof ofthe 4-2 coal seam will collapse the survey line 1 willproduce displacement and the overlying strata in the goafwill have a large overhang distance When the workingface is advanced to 100m the key strata (22m-thicksiltstone layer) fracture and collapse the survey lines 1and 2 produce displacement and the subsidence trend isconsistent +e working face continues to advance to130m 160m 190m and 210m and the rock layer abovethe key strata begins to move and runs through the goafleft by the mining of the 2-2 coal seam +e subsidencetrend of survey lines 1 and 2 below the key strata isconsistent while that of survey lines 3 4 and 5 abovethe key layer is consistent +is indicates that after the keystratum is fractured the soft strata under its control sinksynchronously and harmonically

According to the similar simulation results hydraulicsupports at the LTCC working face in the goaf were subjectedto high loads aggravated by high dynamic load factors +eoverlying stratum exhibited a cutting-type breaking Afterbreaking blocks hinged to each other forming a structure with

Stress compensation 005MPa

Measuringline 5

Measuringline 4

Measuringline 3

Measuringline 2

Measuringline 1

3000mm

Sandy mudstone

Sandy mudstone

Sandy mudstone

Siltstone

Key stratum

2-2 coal seam

4-2 coal seam Loading device

Figure 5 +e similar physical model

6 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

distance During the advancing of the working face a certainroof suspension occurred due to excessively large tensilestrength of the key stratum When the ultimate caving

interval was reached in the key stratum the latter broke andthere was the synergistic breaking of the soft rocks above itAfter breaking the rock strata load was transferred to

Table 1 Lithologic features

Rock formation Lithology +ickness (m) Lithologic characteristics

Overburden

Siltstone 80 Gray silty structure semihard homogeneous bedding and quartz-based compositionSandy mudstone 120 Gray argillaceous structure and semihard

2-2 coal 35 Black blocky structure weak bituminous sheen and small amount of pyriteSandy mudstone 190 Gray argillaceous structure semihard and coal-bearing wire

Key stratum Siltstone 220 Gray silty structure hard homogeneous bedding and quartz-based compositionOverburden Sandy mudstone 240 Gray argillaceous structure and semihardCoal seam 4-2 coal 65 Black blocky structure and weak bituminous sheenFloor Sandy mudstone 50 Gray argillaceous structure semihard and coal-bearing wire

Study area (buertai mine)

Overburden caving

Key stratum2-2 coal seam goaf

4-2 coal seam

Inner mongolia autonomous region

China

Ordos

Figure 1 Mine location and working face layout

20

40

60

Wor

king

resis

tanc

e (M

Pa)

Advance distance (m)

100

120130

80

40

20

60

01

25Support number50

75100

125 150

Ground pressure behavior 64MPa

60MPa

56MPa

52MPa

48MPa

44MPa

40MPa

36MPa

32MPa

28MPa

24MPa

20MPa

16MPa

Figure 2 Morphology of borehole walls at different heights

Shock and Vibration 3

supports and coal bodies +e larger the caving interval ofthe key stratum the larger the load acting on the supports

According to the engineering and geological conditionsof the working face 42107 the key stratum thickness was22m its tensile strength was 54MPa and the load exertedby the overlying stratum on the key stratum was 10MPaUsing formula (4) the cyclic breaking interval of the keystratum was assessed as 295m

4 Similar Physical Simulation Test

41 5e Similar Physical Model +e similar physical modeladopted the burial depth thickness and mining-to-cavingratio of the coal seam 4-2 of 460m 65m and 1 086respectively In situ coal samples were collected by coredrilling and used to fabricate cylindrical specimens formechanical tests Considering the discontinuity of rockmaterials the joint length and continuity factor were set to07 +e mechanical parameters of the coal seam 4-2 roofand floor rocks are listed in Table 2

+e geometric similarity ratioCD between the simulationand field-measured parameters was set to 1 100 +e sim-ilarity ratios of the volumetric weights of rocks and coal wereCc-Rock 1 156 and Cc-Coal 1 1 respectively Based on alarge number of tests and simulations the ground pressuresimilarity theory was developed [24 25] which implies thefollowing similarity ratios of model stresses elastic moduliloads and time

According to the similarity theory the following rela-tionship exists between the physical model and geologicalprototype

Cσ CcCD (5)

where σ is stress and Cσ is the stress similarity ratio c is thebulk density Cc is the bulk density similarity ratio D isdefined as length and CD is the geometric similarity ratio

In addition

CD CδCε (6)

where δ is the displacement Cδ is the displacement similarityratio ε is the strain and Cε is the strain similarity ratio

Finally

CE CσCε (7)

where E is the elastic modulus and CE is the elastic modulussimilarity ratio

Similar physical simulations also use such parameters astensile and compressive strength values (σt and σc) cohesiveforce (c) internal friction angle (β) Poissonrsquos ratio (v) andfriction coefficient (μ)

Insofar as parameters of the physical model are equal tothose of the geological prototype the similarity ratios ofstrain internal friction angle friction coefficient andPoissonrsquos ratio are equal to unity

Cε 1

Cβ 1

Cμ 1

C] 1

(8)

Formulas (5)ndash(8) yield the following equality

Cσ CE Cc Cσc Cσt

(9)

For the above values of Cc-Rock 1 156 and Cc-Coal 1 1we get the following stress and elastic modulus similarityratios

Cσ CE CcCD 156 (10)

Rib heave

Floor heave

(a)

Rib spalling

(b)

Figure 3 Strong ground pressure manifestations

b

h

q

M

XY

Z

ymax

Figure 4 Mechanical model of a cantilever beam pure bending

4 Shock and Vibration

+en the load similarity ratio is

CF CcCD3 156 times 106 (11)

+e time similarity ratio is

Ct CD

1113968 10 (12)

Using the obtained relation between stress and elasticmodulus CσE 1 156 the physical-mechanical parametersof the stratum under similar simulation were determinedand the initial mix ratios of similar materials were deter-mined from the similar theory [26 27] Materials used forsimilar simulations included fine sand with the particlediameter range of 025ndash035mm lime and gypsum In orderto obtain the appropriate elastic modulus of the materialfor similar simulation a large number of mechanical testswere performed for the collected specimens +e mixratios of rock similar materials are listed in the lastcolumn of Table 1 First the initial mix ratios of similarmaterials were preset +en weighted mixing was per-formed for similar materials (fine sand lime and gyp-sum) which were used to prepare standard columnarspecimens After curing for seven days these specimenswere subjected to mechanical tests +e final mix ratios ofsimilar materials were determined after screening andcomparative analysis

+e specification of the similar simulation device was asfollows length times width times height 3000 mm times 400 mm times

2000mm Fine sand was used as aggregate lime and gypsumwere used as a cementing agent and water was used as abonding agent Based on the enrichment status of theworking face 42107 in the Buertai Coal Mine the modelspecificationwas as follows lengthtimes widthtimes height 3000mmtimes 400mmtimes 1500mm +e laying thickness of the coal seam2-2 was 35mm and that of the overlying stratum of thecoal seam 2-2 was 700 mm +e laying thickness of thecoal seam 4-2 was 65m and that of the stratum betweencoal seams 2-2 and 4-2 was 650mm the layingthickness of the floor of the coal seam 4-2 was 50mm In

the simulation experiment the advancing distance alongthe orientation of the working face was 2200 mm thedistances from the model left and right boundaries to thesetup entry and coal seam stop mining line were bothequal to 400 mm To achieve uniform loading a multi-function adjustable physical simulation loading devicewas installed above the testing device which partiallycompensated the gravity-induced load of the overlyingstratum by the imposed load 005MPa +ere are 5 mea-suring lines of displacement arranged in the physical model+e physical model is depicted in Figure 5

+e time similarity ratio was the ratio of time periodsrequired for excavations in the physical model and the actualoperation at the working face Since the roof movement isconsiderably affected by the extraction speed of the workingface the ground pressure manifestation pattern shouldchange dramatically with the extraction speed variation atthe working +erefore the time similarity ratio was used todetermine the time needed for excavation in the simulationmodel Based on the time and geometric similarity ratioseach day of excavation at the real working face correspondsto 24 h of excavation in the physical model+e 38 workingschedule is adopted in the Buertai Coal Mine with two shiftsper day To be specific every shift is responsible for exca-vation for 50mm in the physical model (excavation for 5min the actual working face) with two shifts per day +at isthe working face in the model advanced by 100m per 24 hwhile that at the actual working face advanced by 10m per24 h

42 Breaking Characteristics of Overburden After the ex-cavation of the coal seam 2-2 was completed its overlyingstratum broke and vertical cracks formed on both sides ofthe goaf +ere was general bending subsidence of theoverlying stratum Due to the mining-induced influence ofthe coal seam 2-2 the overall strength of the overlyingstratum decreased and the physical-mechanical propertiesof rocks became close to those of soft rocks as shown in

Table 2 Physical-mechanical parameters of coal rocks and mix ratios of rock similar materials

NameStrata

thickness(m)

Tensilestrength(MPa)

Elasticmodulus(GPa)

Cohesion(MPa)

Compressive strength Bulk density Mix ratio

Prototype(MPa)

Model(MPa)

Prototype(kgm3)

Model(kgm3)

Sand gypsum lime

Sandymudstone 500 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 80 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 120 157 634 36 3263 021 2100 1350 7 7 3

2-2 coal 35 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 190 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 220 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 240 157 634 36 3263 021 2100 1350 7 7 3

4-2 coal 65 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 50 157 634 36 3263 021 2100 1350 7 7 3

Shock and Vibration 5

Figure 6(a) After the breaking of the key stratum betweencoal seams 2-2 and 4-2 there would be a synergisticcollapse of the overlying stratum of coal seam 2-2(equivalent to soft rocks) and key stratum When theworking face 42107 advanced to 85m the lower immediateroof began to fall along with the excavation while theimmediate upper roof and key stratum remained intact Atthis moment the breaking angle of the lower immediate roofwas about 52deg as shown in Figure 6(b)

As the working face continued to advance the de-formation of the immediate upper roof and key stratumgradually increased and cracks began to develop Whenthe working face advanced to 100m the 22m-thick keystratum broke in the rear upper part of the coal wall alongthe working face and the breaking angle was about 60deg asshown in Figure 6(c) At this time the working face wassubject to the first weighting which was consistent with anincrease in the support load when the working face ad-vanced to 100m in Figure 2

+e bending deformation of the key stratum and bedseparation space below it began to increase with the workingface advance First transverse cracks appeared in the keystratum +en vertical cracks appeared in the middle of thebed separation space exhibiting a bottom-up developmentWhen the working face advanced to 130m vertical crackspenetrated the entire key stratum +e key stratum brokeand as a result the immediate roof key stratum and load-bearing layer underwent the first cutting as a whole in therear upper part of the coal wall along the working face +ecutting angle was about 64deg as shown in Figure 6(d)Meanwhile there was a synergistic movement of the keystratum and the overlying stratum of the coal seam 2-2where the extraction was completed (equivalent to softrocks)

As the working face continued to advance after theextraction of coal seam 2-2 was completed (equivalent tosoft rocks) the key load-bearing and overlying strataexhibited a cyclic synergistic cutting-type breaking +ecutting position was found in the rear upper part of the coal

wall along the working face +e cyclic cutting angle wasclose to the initial cutting angle as shown in Figures 6(e) and6(f ) +e overlying stratum underwent a cyclic cutting-typebreaking as a whole Two adjacent cutting blocks were inclose contact with each other squeezing each other andforming a hinged structure+is structure subsided slowly asthe overlying stratum broke until it came into full contactwith gangues in the goaf

43 Displacement Characteristics of Overburden Figure 7shows the overburden displacement curve in the miningprocess of number 4-2 coal seam It can be seen fromFigure 7 that the overlying strata of number 4-2 coal seamexperienced a continuous dynamic subsidence andmovement process during the mining process +e sub-sidence trend of the overlying strata is nonlinear and themovement form is asymmetric Specifically when theworking face is advanced to 85m the immediate roof ofthe 4-2 coal seam will collapse the survey line 1 willproduce displacement and the overlying strata in the goafwill have a large overhang distance When the workingface is advanced to 100m the key strata (22m-thicksiltstone layer) fracture and collapse the survey lines 1and 2 produce displacement and the subsidence trend isconsistent +e working face continues to advance to130m 160m 190m and 210m and the rock layer abovethe key strata begins to move and runs through the goafleft by the mining of the 2-2 coal seam +e subsidencetrend of survey lines 1 and 2 below the key strata isconsistent while that of survey lines 3 4 and 5 abovethe key layer is consistent +is indicates that after the keystratum is fractured the soft strata under its control sinksynchronously and harmonically

According to the similar simulation results hydraulicsupports at the LTCC working face in the goaf were subjectedto high loads aggravated by high dynamic load factors +eoverlying stratum exhibited a cutting-type breaking Afterbreaking blocks hinged to each other forming a structure with

Stress compensation 005MPa

Measuringline 5

Measuringline 4

Measuringline 3

Measuringline 2

Measuringline 1

3000mm

Sandy mudstone

Sandy mudstone

Sandy mudstone

Siltstone

Key stratum

2-2 coal seam

4-2 coal seam Loading device

Figure 5 +e similar physical model

6 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

supports and coal bodies +e larger the caving interval ofthe key stratum the larger the load acting on the supports

According to the engineering and geological conditionsof the working face 42107 the key stratum thickness was22m its tensile strength was 54MPa and the load exertedby the overlying stratum on the key stratum was 10MPaUsing formula (4) the cyclic breaking interval of the keystratum was assessed as 295m

4 Similar Physical Simulation Test

41 5e Similar Physical Model +e similar physical modeladopted the burial depth thickness and mining-to-cavingratio of the coal seam 4-2 of 460m 65m and 1 086respectively In situ coal samples were collected by coredrilling and used to fabricate cylindrical specimens formechanical tests Considering the discontinuity of rockmaterials the joint length and continuity factor were set to07 +e mechanical parameters of the coal seam 4-2 roofand floor rocks are listed in Table 2

+e geometric similarity ratioCD between the simulationand field-measured parameters was set to 1 100 +e sim-ilarity ratios of the volumetric weights of rocks and coal wereCc-Rock 1 156 and Cc-Coal 1 1 respectively Based on alarge number of tests and simulations the ground pressuresimilarity theory was developed [24 25] which implies thefollowing similarity ratios of model stresses elastic moduliloads and time

According to the similarity theory the following rela-tionship exists between the physical model and geologicalprototype

Cσ CcCD (5)

where σ is stress and Cσ is the stress similarity ratio c is thebulk density Cc is the bulk density similarity ratio D isdefined as length and CD is the geometric similarity ratio

In addition

CD CδCε (6)

where δ is the displacement Cδ is the displacement similarityratio ε is the strain and Cε is the strain similarity ratio

Finally

CE CσCε (7)

where E is the elastic modulus and CE is the elastic modulussimilarity ratio

Similar physical simulations also use such parameters astensile and compressive strength values (σt and σc) cohesiveforce (c) internal friction angle (β) Poissonrsquos ratio (v) andfriction coefficient (μ)

Insofar as parameters of the physical model are equal tothose of the geological prototype the similarity ratios ofstrain internal friction angle friction coefficient andPoissonrsquos ratio are equal to unity

Cε 1

Cβ 1

Cμ 1

C] 1

(8)

Formulas (5)ndash(8) yield the following equality

Cσ CE Cc Cσc Cσt

(9)

For the above values of Cc-Rock 1 156 and Cc-Coal 1 1we get the following stress and elastic modulus similarityratios

Cσ CE CcCD 156 (10)

Rib heave

Floor heave

(a)

Rib spalling

(b)

Figure 3 Strong ground pressure manifestations

b

h

q

M

XY

Z

ymax

Figure 4 Mechanical model of a cantilever beam pure bending

4 Shock and Vibration

+en the load similarity ratio is

CF CcCD3 156 times 106 (11)

+e time similarity ratio is

Ct CD

1113968 10 (12)

Using the obtained relation between stress and elasticmodulus CσE 1 156 the physical-mechanical parametersof the stratum under similar simulation were determinedand the initial mix ratios of similar materials were deter-mined from the similar theory [26 27] Materials used forsimilar simulations included fine sand with the particlediameter range of 025ndash035mm lime and gypsum In orderto obtain the appropriate elastic modulus of the materialfor similar simulation a large number of mechanical testswere performed for the collected specimens +e mixratios of rock similar materials are listed in the lastcolumn of Table 1 First the initial mix ratios of similarmaterials were preset +en weighted mixing was per-formed for similar materials (fine sand lime and gyp-sum) which were used to prepare standard columnarspecimens After curing for seven days these specimenswere subjected to mechanical tests +e final mix ratios ofsimilar materials were determined after screening andcomparative analysis

+e specification of the similar simulation device was asfollows length times width times height 3000 mm times 400 mm times

2000mm Fine sand was used as aggregate lime and gypsumwere used as a cementing agent and water was used as abonding agent Based on the enrichment status of theworking face 42107 in the Buertai Coal Mine the modelspecificationwas as follows lengthtimes widthtimes height 3000mmtimes 400mmtimes 1500mm +e laying thickness of the coal seam2-2 was 35mm and that of the overlying stratum of thecoal seam 2-2 was 700 mm +e laying thickness of thecoal seam 4-2 was 65m and that of the stratum betweencoal seams 2-2 and 4-2 was 650mm the layingthickness of the floor of the coal seam 4-2 was 50mm In

the simulation experiment the advancing distance alongthe orientation of the working face was 2200 mm thedistances from the model left and right boundaries to thesetup entry and coal seam stop mining line were bothequal to 400 mm To achieve uniform loading a multi-function adjustable physical simulation loading devicewas installed above the testing device which partiallycompensated the gravity-induced load of the overlyingstratum by the imposed load 005MPa +ere are 5 mea-suring lines of displacement arranged in the physical model+e physical model is depicted in Figure 5

+e time similarity ratio was the ratio of time periodsrequired for excavations in the physical model and the actualoperation at the working face Since the roof movement isconsiderably affected by the extraction speed of the workingface the ground pressure manifestation pattern shouldchange dramatically with the extraction speed variation atthe working +erefore the time similarity ratio was used todetermine the time needed for excavation in the simulationmodel Based on the time and geometric similarity ratioseach day of excavation at the real working face correspondsto 24 h of excavation in the physical model+e 38 workingschedule is adopted in the Buertai Coal Mine with two shiftsper day To be specific every shift is responsible for exca-vation for 50mm in the physical model (excavation for 5min the actual working face) with two shifts per day +at isthe working face in the model advanced by 100m per 24 hwhile that at the actual working face advanced by 10m per24 h

42 Breaking Characteristics of Overburden After the ex-cavation of the coal seam 2-2 was completed its overlyingstratum broke and vertical cracks formed on both sides ofthe goaf +ere was general bending subsidence of theoverlying stratum Due to the mining-induced influence ofthe coal seam 2-2 the overall strength of the overlyingstratum decreased and the physical-mechanical propertiesof rocks became close to those of soft rocks as shown in

Table 2 Physical-mechanical parameters of coal rocks and mix ratios of rock similar materials

NameStrata

thickness(m)

Tensilestrength(MPa)

Elasticmodulus(GPa)

Cohesion(MPa)

Compressive strength Bulk density Mix ratio

Prototype(MPa)

Model(MPa)

Prototype(kgm3)

Model(kgm3)

Sand gypsum lime

Sandymudstone 500 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 80 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 120 157 634 36 3263 021 2100 1350 7 7 3

2-2 coal 35 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 190 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 220 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 240 157 634 36 3263 021 2100 1350 7 7 3

4-2 coal 65 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 50 157 634 36 3263 021 2100 1350 7 7 3

Shock and Vibration 5

Figure 6(a) After the breaking of the key stratum betweencoal seams 2-2 and 4-2 there would be a synergisticcollapse of the overlying stratum of coal seam 2-2(equivalent to soft rocks) and key stratum When theworking face 42107 advanced to 85m the lower immediateroof began to fall along with the excavation while theimmediate upper roof and key stratum remained intact Atthis moment the breaking angle of the lower immediate roofwas about 52deg as shown in Figure 6(b)

As the working face continued to advance the de-formation of the immediate upper roof and key stratumgradually increased and cracks began to develop Whenthe working face advanced to 100m the 22m-thick keystratum broke in the rear upper part of the coal wall alongthe working face and the breaking angle was about 60deg asshown in Figure 6(c) At this time the working face wassubject to the first weighting which was consistent with anincrease in the support load when the working face ad-vanced to 100m in Figure 2

+e bending deformation of the key stratum and bedseparation space below it began to increase with the workingface advance First transverse cracks appeared in the keystratum +en vertical cracks appeared in the middle of thebed separation space exhibiting a bottom-up developmentWhen the working face advanced to 130m vertical crackspenetrated the entire key stratum +e key stratum brokeand as a result the immediate roof key stratum and load-bearing layer underwent the first cutting as a whole in therear upper part of the coal wall along the working face +ecutting angle was about 64deg as shown in Figure 6(d)Meanwhile there was a synergistic movement of the keystratum and the overlying stratum of the coal seam 2-2where the extraction was completed (equivalent to softrocks)

As the working face continued to advance after theextraction of coal seam 2-2 was completed (equivalent tosoft rocks) the key load-bearing and overlying strataexhibited a cyclic synergistic cutting-type breaking +ecutting position was found in the rear upper part of the coal

wall along the working face +e cyclic cutting angle wasclose to the initial cutting angle as shown in Figures 6(e) and6(f ) +e overlying stratum underwent a cyclic cutting-typebreaking as a whole Two adjacent cutting blocks were inclose contact with each other squeezing each other andforming a hinged structure+is structure subsided slowly asthe overlying stratum broke until it came into full contactwith gangues in the goaf

43 Displacement Characteristics of Overburden Figure 7shows the overburden displacement curve in the miningprocess of number 4-2 coal seam It can be seen fromFigure 7 that the overlying strata of number 4-2 coal seamexperienced a continuous dynamic subsidence andmovement process during the mining process +e sub-sidence trend of the overlying strata is nonlinear and themovement form is asymmetric Specifically when theworking face is advanced to 85m the immediate roof ofthe 4-2 coal seam will collapse the survey line 1 willproduce displacement and the overlying strata in the goafwill have a large overhang distance When the workingface is advanced to 100m the key strata (22m-thicksiltstone layer) fracture and collapse the survey lines 1and 2 produce displacement and the subsidence trend isconsistent +e working face continues to advance to130m 160m 190m and 210m and the rock layer abovethe key strata begins to move and runs through the goafleft by the mining of the 2-2 coal seam +e subsidencetrend of survey lines 1 and 2 below the key strata isconsistent while that of survey lines 3 4 and 5 abovethe key layer is consistent +is indicates that after the keystratum is fractured the soft strata under its control sinksynchronously and harmonically

According to the similar simulation results hydraulicsupports at the LTCC working face in the goaf were subjectedto high loads aggravated by high dynamic load factors +eoverlying stratum exhibited a cutting-type breaking Afterbreaking blocks hinged to each other forming a structure with

Stress compensation 005MPa

Measuringline 5

Measuringline 4

Measuringline 3

Measuringline 2

Measuringline 1

3000mm

Sandy mudstone

Sandy mudstone

Sandy mudstone

Siltstone

Key stratum

2-2 coal seam

4-2 coal seam Loading device

Figure 5 +e similar physical model

6 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

+en the load similarity ratio is

CF CcCD3 156 times 106 (11)

+e time similarity ratio is

Ct CD

1113968 10 (12)

Using the obtained relation between stress and elasticmodulus CσE 1 156 the physical-mechanical parametersof the stratum under similar simulation were determinedand the initial mix ratios of similar materials were deter-mined from the similar theory [26 27] Materials used forsimilar simulations included fine sand with the particlediameter range of 025ndash035mm lime and gypsum In orderto obtain the appropriate elastic modulus of the materialfor similar simulation a large number of mechanical testswere performed for the collected specimens +e mixratios of rock similar materials are listed in the lastcolumn of Table 1 First the initial mix ratios of similarmaterials were preset +en weighted mixing was per-formed for similar materials (fine sand lime and gyp-sum) which were used to prepare standard columnarspecimens After curing for seven days these specimenswere subjected to mechanical tests +e final mix ratios ofsimilar materials were determined after screening andcomparative analysis

+e specification of the similar simulation device was asfollows length times width times height 3000 mm times 400 mm times

2000mm Fine sand was used as aggregate lime and gypsumwere used as a cementing agent and water was used as abonding agent Based on the enrichment status of theworking face 42107 in the Buertai Coal Mine the modelspecificationwas as follows lengthtimes widthtimes height 3000mmtimes 400mmtimes 1500mm +e laying thickness of the coal seam2-2 was 35mm and that of the overlying stratum of thecoal seam 2-2 was 700 mm +e laying thickness of thecoal seam 4-2 was 65m and that of the stratum betweencoal seams 2-2 and 4-2 was 650mm the layingthickness of the floor of the coal seam 4-2 was 50mm In

the simulation experiment the advancing distance alongthe orientation of the working face was 2200 mm thedistances from the model left and right boundaries to thesetup entry and coal seam stop mining line were bothequal to 400 mm To achieve uniform loading a multi-function adjustable physical simulation loading devicewas installed above the testing device which partiallycompensated the gravity-induced load of the overlyingstratum by the imposed load 005MPa +ere are 5 mea-suring lines of displacement arranged in the physical model+e physical model is depicted in Figure 5

+e time similarity ratio was the ratio of time periodsrequired for excavations in the physical model and the actualoperation at the working face Since the roof movement isconsiderably affected by the extraction speed of the workingface the ground pressure manifestation pattern shouldchange dramatically with the extraction speed variation atthe working +erefore the time similarity ratio was used todetermine the time needed for excavation in the simulationmodel Based on the time and geometric similarity ratioseach day of excavation at the real working face correspondsto 24 h of excavation in the physical model+e 38 workingschedule is adopted in the Buertai Coal Mine with two shiftsper day To be specific every shift is responsible for exca-vation for 50mm in the physical model (excavation for 5min the actual working face) with two shifts per day +at isthe working face in the model advanced by 100m per 24 hwhile that at the actual working face advanced by 10m per24 h

42 Breaking Characteristics of Overburden After the ex-cavation of the coal seam 2-2 was completed its overlyingstratum broke and vertical cracks formed on both sides ofthe goaf +ere was general bending subsidence of theoverlying stratum Due to the mining-induced influence ofthe coal seam 2-2 the overall strength of the overlyingstratum decreased and the physical-mechanical propertiesof rocks became close to those of soft rocks as shown in

Table 2 Physical-mechanical parameters of coal rocks and mix ratios of rock similar materials

NameStrata

thickness(m)

Tensilestrength(MPa)

Elasticmodulus(GPa)

Cohesion(MPa)

Compressive strength Bulk density Mix ratio

Prototype(MPa)

Model(MPa)

Prototype(kgm3)

Model(kgm3)

Sand gypsum lime

Sandymudstone 500 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 80 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 120 157 634 36 3263 021 2100 1350 7 7 3

2-2 coal 35 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 190 157 634 36 3263 021 2100 1350 7 7 3

Siltstone 220 431 1052 54 8634 055 2550 1640 9 5 5Sandymudstone 240 157 634 36 3263 021 2100 1350 7 7 3

4-2 coal 65 125 544 28 1078 007 1300 1300 10 5 5Sandymudstone 50 157 634 36 3263 021 2100 1350 7 7 3

Shock and Vibration 5

Figure 6(a) After the breaking of the key stratum betweencoal seams 2-2 and 4-2 there would be a synergisticcollapse of the overlying stratum of coal seam 2-2(equivalent to soft rocks) and key stratum When theworking face 42107 advanced to 85m the lower immediateroof began to fall along with the excavation while theimmediate upper roof and key stratum remained intact Atthis moment the breaking angle of the lower immediate roofwas about 52deg as shown in Figure 6(b)

As the working face continued to advance the de-formation of the immediate upper roof and key stratumgradually increased and cracks began to develop Whenthe working face advanced to 100m the 22m-thick keystratum broke in the rear upper part of the coal wall alongthe working face and the breaking angle was about 60deg asshown in Figure 6(c) At this time the working face wassubject to the first weighting which was consistent with anincrease in the support load when the working face ad-vanced to 100m in Figure 2

+e bending deformation of the key stratum and bedseparation space below it began to increase with the workingface advance First transverse cracks appeared in the keystratum +en vertical cracks appeared in the middle of thebed separation space exhibiting a bottom-up developmentWhen the working face advanced to 130m vertical crackspenetrated the entire key stratum +e key stratum brokeand as a result the immediate roof key stratum and load-bearing layer underwent the first cutting as a whole in therear upper part of the coal wall along the working face +ecutting angle was about 64deg as shown in Figure 6(d)Meanwhile there was a synergistic movement of the keystratum and the overlying stratum of the coal seam 2-2where the extraction was completed (equivalent to softrocks)

As the working face continued to advance after theextraction of coal seam 2-2 was completed (equivalent tosoft rocks) the key load-bearing and overlying strataexhibited a cyclic synergistic cutting-type breaking +ecutting position was found in the rear upper part of the coal

wall along the working face +e cyclic cutting angle wasclose to the initial cutting angle as shown in Figures 6(e) and6(f ) +e overlying stratum underwent a cyclic cutting-typebreaking as a whole Two adjacent cutting blocks were inclose contact with each other squeezing each other andforming a hinged structure+is structure subsided slowly asthe overlying stratum broke until it came into full contactwith gangues in the goaf

43 Displacement Characteristics of Overburden Figure 7shows the overburden displacement curve in the miningprocess of number 4-2 coal seam It can be seen fromFigure 7 that the overlying strata of number 4-2 coal seamexperienced a continuous dynamic subsidence andmovement process during the mining process +e sub-sidence trend of the overlying strata is nonlinear and themovement form is asymmetric Specifically when theworking face is advanced to 85m the immediate roof ofthe 4-2 coal seam will collapse the survey line 1 willproduce displacement and the overlying strata in the goafwill have a large overhang distance When the workingface is advanced to 100m the key strata (22m-thicksiltstone layer) fracture and collapse the survey lines 1and 2 produce displacement and the subsidence trend isconsistent +e working face continues to advance to130m 160m 190m and 210m and the rock layer abovethe key strata begins to move and runs through the goafleft by the mining of the 2-2 coal seam +e subsidencetrend of survey lines 1 and 2 below the key strata isconsistent while that of survey lines 3 4 and 5 abovethe key layer is consistent +is indicates that after the keystratum is fractured the soft strata under its control sinksynchronously and harmonically

According to the similar simulation results hydraulicsupports at the LTCC working face in the goaf were subjectedto high loads aggravated by high dynamic load factors +eoverlying stratum exhibited a cutting-type breaking Afterbreaking blocks hinged to each other forming a structure with

Stress compensation 005MPa

Measuringline 5

Measuringline 4

Measuringline 3

Measuringline 2

Measuringline 1

3000mm

Sandy mudstone

Sandy mudstone

Sandy mudstone

Siltstone

Key stratum

2-2 coal seam

4-2 coal seam Loading device

Figure 5 +e similar physical model

6 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

Figure 6(a) After the breaking of the key stratum betweencoal seams 2-2 and 4-2 there would be a synergisticcollapse of the overlying stratum of coal seam 2-2(equivalent to soft rocks) and key stratum When theworking face 42107 advanced to 85m the lower immediateroof began to fall along with the excavation while theimmediate upper roof and key stratum remained intact Atthis moment the breaking angle of the lower immediate roofwas about 52deg as shown in Figure 6(b)

As the working face continued to advance the de-formation of the immediate upper roof and key stratumgradually increased and cracks began to develop Whenthe working face advanced to 100m the 22m-thick keystratum broke in the rear upper part of the coal wall alongthe working face and the breaking angle was about 60deg asshown in Figure 6(c) At this time the working face wassubject to the first weighting which was consistent with anincrease in the support load when the working face ad-vanced to 100m in Figure 2

+e bending deformation of the key stratum and bedseparation space below it began to increase with the workingface advance First transverse cracks appeared in the keystratum +en vertical cracks appeared in the middle of thebed separation space exhibiting a bottom-up developmentWhen the working face advanced to 130m vertical crackspenetrated the entire key stratum +e key stratum brokeand as a result the immediate roof key stratum and load-bearing layer underwent the first cutting as a whole in therear upper part of the coal wall along the working face +ecutting angle was about 64deg as shown in Figure 6(d)Meanwhile there was a synergistic movement of the keystratum and the overlying stratum of the coal seam 2-2where the extraction was completed (equivalent to softrocks)

As the working face continued to advance after theextraction of coal seam 2-2 was completed (equivalent tosoft rocks) the key load-bearing and overlying strataexhibited a cyclic synergistic cutting-type breaking +ecutting position was found in the rear upper part of the coal

wall along the working face +e cyclic cutting angle wasclose to the initial cutting angle as shown in Figures 6(e) and6(f ) +e overlying stratum underwent a cyclic cutting-typebreaking as a whole Two adjacent cutting blocks were inclose contact with each other squeezing each other andforming a hinged structure+is structure subsided slowly asthe overlying stratum broke until it came into full contactwith gangues in the goaf

43 Displacement Characteristics of Overburden Figure 7shows the overburden displacement curve in the miningprocess of number 4-2 coal seam It can be seen fromFigure 7 that the overlying strata of number 4-2 coal seamexperienced a continuous dynamic subsidence andmovement process during the mining process +e sub-sidence trend of the overlying strata is nonlinear and themovement form is asymmetric Specifically when theworking face is advanced to 85m the immediate roof ofthe 4-2 coal seam will collapse the survey line 1 willproduce displacement and the overlying strata in the goafwill have a large overhang distance When the workingface is advanced to 100m the key strata (22m-thicksiltstone layer) fracture and collapse the survey lines 1and 2 produce displacement and the subsidence trend isconsistent +e working face continues to advance to130m 160m 190m and 210m and the rock layer abovethe key strata begins to move and runs through the goafleft by the mining of the 2-2 coal seam +e subsidencetrend of survey lines 1 and 2 below the key strata isconsistent while that of survey lines 3 4 and 5 abovethe key layer is consistent +is indicates that after the keystratum is fractured the soft strata under its control sinksynchronously and harmonically

According to the similar simulation results hydraulicsupports at the LTCC working face in the goaf were subjectedto high loads aggravated by high dynamic load factors +eoverlying stratum exhibited a cutting-type breaking Afterbreaking blocks hinged to each other forming a structure with

Stress compensation 005MPa

Measuringline 5

Measuringline 4

Measuringline 3

Measuringline 2

Measuringline 1

3000mm

Sandy mudstone

Sandy mudstone

Sandy mudstone

Siltstone

Key stratum

2-2 coal seam

4-2 coal seam Loading device

Figure 5 +e similar physical model

6 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

a dominated sliding instability Hereinafter such structures arereferred to as cutting blocks as shown in Figure 8

During the advancing of the working face the cutting blockinterval is influenced by the key stratum breaking interval Asseen in Figure 8 the first cutting block interval (equivalent tothat of the key stratum) occurred at the working face advanceof 130m At this time supports of the working face withstoodthe load of the entire cutting block +e support workingresistance rose abruptly which was consistent with the supportload increase at the working face advance of 130m in Figure 2At this time a strong ground pressure began to manifest itselfat the working faceWhen the working face advanced to 160mthe second cyclic breaking of the cutting blocks occurred +ecaving interval of cutting blocks (equivalent to that of the keystratum) was 30m so the similar simulation results were veryclose to those predicted by themechanical model (295m)+isclose correlation corroborated similar simulation accuracy

44 Analysis of Support Working Resistance +e stressanalysis was performed for cutting block A depicted in Fig-ure 8 using the geometric scheme presented in Figure 9 [28]

HereQ is the compressive stress exerted by the overlyingstratum of the coal seam 2-2 (equivalent to soft rocks) onthe cutting block kN T is the horizontal squeeze forcebetween the cutting blocks kN G is the dead weight of thecutting block kN F is friction force acting between cuttingblocks kN p is support load kN a is the cutting blockheight m L is the cutting interval m x is the distancebetween the application point of the support and coal wallm α is the cutting angle degree +e support working re-sistance was calculated for the cutting block structure usingthe rotational equilibrium principle

QL

2+ a cot α1113874 1113875 + G

L

2+

a

2cot α1113874 1113875 minus px minus FL sin α minus TOD 0

TOD Ta

2sin α + T L +

a

2cot α1113874 1113875cos α

⎧⎪⎪⎪⎪⎪⎨

⎪⎪⎪⎪⎪⎩

(13)

where F fT f is the friction coefficient between blocks andT ξG ξ is the horizontal squeeze factor

+e support load p can be derived via formula (13) asfollows

Overburden caving

Longitudinal cracks

(a)

Key stratum

Immediate roof caving52deg

(b)

Key stratum caving

60deg

(c)

Cutting block

Overburden caving again

Cantilever beam caving

64deg

(d)

Cutting block

(e)

Key stratum caving

Cutting block

Overburden caving again

(f )

Figure 6 Overlying stratum breaking and instability in similar simulations (a) extraction of the 2-2 coal seam completed (b) extraction ofthe 4-2 coal seam at 85m (c) extraction of the 4-2 coal seam at 100m (d) extraction of the 4-2 coal seam at 130m (e) extraction of the 4-2 coal seam at 160m (f ) completed extraction of the 4-2 coal seam

Shock and Vibration 7

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

25 50 75 100 125Measuring points (cm)

1

(a)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

30 60 90 120 150Measuring points (cm)

12

(b)

123

45

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180Measuring points (cm)

(c)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0Su

bsid

ence

(m)

30 60 90 120 150 180 210Measuring points (cm)

123

45

(d)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

20 40 60 80 100 120 140 160 180 200 220 240Measuring points (cm)

123

45

(e)

ndash6

ndash5

ndash4

ndash3

ndash2

ndash1

0

Subs

iden

ce (m

)

40 80 120 160 200 240 280Measuring points (cm)

123

45

(f )

Figure 7 Overlying stratum subsidence in similar simulations (a) extraction of the 4-2 coal seam at 85m (b) extraction of the 4-2 coalseam at 100m (c) extraction of the 4-2 coal seam at 130m (d) extraction of the 4-2 coal seam at 160m (e) extraction of the 4-2 coal seamat 190m (f ) completed extraction of the 4-2 coal seam

8 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

p L(Q + G minus 2ξfG sin α minus 2ξG cos α) + a(2Q cot α + G cot α minus ξG sin α minus ξG cos α cot α)

2x (14)

+e width of the working face 42107 was 300m and thecoal seam thickness was 65m +e support model wasZFY210002539D +e thickness of the cutting block was65m and the bulk density was 25 kNm3 +e average cyclicweighting interval was 30m +e distance from the appli-cation point of the resultant force of support to the coal wallwas 45m+e friction factor between the rock blocks was setto 07 the horizontal squeeze factor was 03 and the cuttingangle was 64deg +e support working resistance via formula(14) was calculated as p 24000 kN which corresponded tothe internal stress of about 52MPa (when the diameter of thesupporting column is 054m the working resistance of thesupporting column is 24times106(314times 0272)2 52MPa)

+is finding complies with measured data of ground pres-sure at the working face as shown in Figure 2 Uponweighing the support working resistance towards the cy-cling end was 5ndash52MPa so that the theoretical calculationresults basically agreed with the field measurements

5 Control Countermeasures for the StrongGround Pressure

51 Hydraulic Fracturing Scheme As analyzed above the22m-thick siltstone (key stratum) at a distance of 24m fromthe coal seam 4-2 was the primary stratum controlling thestrong ground pressure manifestation at the working face

AB

2-2coal seam

4-2coal seam

59deg

62deg 68deg

68deg

100m 30m 30m 30m 30m

Cuttingblock 1

Cuttingblock 2

Cuttingblock 3

Cuttingblock 4

Existing cracksPartially closed cracksOpenednew cracks

Figure 8 Structural sketch of the cutting blocks

L

α

a

F

F

OP x

A

BC

G

TD

X

Y

Q

E

Ty

Tx

Figure 9 Stress analysis of cutting blocks

Shock and Vibration 9

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

42107 +erefore the directional hydraulic fracturing wasmainly applied to the key stratum between both coal seamsin order to reduce the caving interval of the key stratum anddiminish the effect of cutting blocks on the ground pressureof the working face

A detailed illustration is provided for the three di-rectional boreholes drilled in the roadway cross section ofworking face 42107 as shown in Figure 10 First thedirectional boreholes K1 SF1 and SF2 were drilledobliquely on the roof of both roadways of the workingface +e direction of the drill bit was constantly adjustedso that the drill bit had just reached the middle of the keystratum when the vertical distance to the roof was 35mm(the distance from the key stratum to the coal seam 4-2was 24m and the key stratum thickness was 22m) +edrilling continued horizontally Borehole SF1 was drilledhorizontally for 140m and the horizontal segment of theborehole was 92m from the haulage roadway of theworking face 42107 Borehole SF2 was drilled horizontallyfor 216m and the horizontal segment of the borehole was142m from the haulage roadway of the working face42107 Borehole K1 was drilled horizontally for 198mand the horizontal segment of the borehole was 150mfrom the ventilation roadway of the working face 42107At the last step hydraulic fracturing was applied to the

22m-thick key stratum to reduce the breaking interval ofthe key stratum

Boreholes SF1 and SF2 were first drilled followed byhydraulic fracturing When the fracturing was over theborehole K1 was drilled and the fracturing was inducedOnly the horizontal segment was drilled for the key stratumto which the hydraulic fracturing was applied +e totallength of three boreholes was 1132m and the total length ofhydraulic fracturing was 554m

52 ApplicationEffect In order to assess the control effect ofhydraulic fracturing of the interseam key stratum on thestrong ground pressure of the working face the data onsupport working resistance within the advancing distance of560ndash950m were used At the working face 42107 a total of125 ZFY210002539D two-column cover-type caving hy-draulic supports were installed Ground pressure dynamicgauges were installed in supports in order to continuouslyrecord the support working resistance at the working face+e measured values of support working resistance versusadvancing distance are plotted in Figure 11

+e opening pressure of the safety valve at the workingface 42107 was 46MPa Before the hydraulic fracturing of the22m-thick key stratum the support working resistance

216m

140m

142m

SF2

Key stratum

198m150m

SF1

92m

Goaf

Headentry

Tailentry

K1

(a)

150mSF2

142m

Tailentry

198mSF1

92mHeadentry

K1

140m

216mSetup room

(b)

Figure 10 Positions and parameters of three boreholes (a) stereogram (b) profile

10 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

upon weighting of the working face 42017 was 50ndash52MPaand the cyclic weighting interval was about 30m Afterhydraulic fracturing of the key stratum the weighting in-terval and intensity of the working face were determined asshown in Figure 11 After the directional hydraulic frac-turing was applied to the interseam key stratum the cyclicweighting interval of the working face 42107 was reduced to169m the continuous length of cyclic weighting was re-duced to 49m and the average support working resistancetowards the cycling end was about 42MPa +e resultsshowed that after hydraulic fracturing has been applied tothe 22m-thick key stratum there was a significant reductionin the cyclic weighting interval continuous length ofweighting and support working resistance towards thecycling end at the working face 42107 +is implies thatstrong ground pressure was efficiently controlled at theworking face

6 Conclusions

+e results obtained made it possible to draw the followingconclusions

(1) Field measurements and physical simulations of theworking face 42107 of the Buertai Coal Mine inChina confirmed strong ground pressure manifes-tations which included rib spalling severe roadwaydeformation high cyclic weighting intensity andhigh dynamic loads +e key stratum between seamsalong with the overlying soft rocks after the ex-traction of the coal seam 2-2 underwent synergisticsliding and breaking

(2) +e structural model for cutting blocks from theoverlying stratum of the working face with the ac-count of the key stratum breaking instability wasproposed After the key stratum collapse there was acutting-type breaking of soft rocks with synergisticdeformation and breaking Finally a hinged struc-ture of cutting blocks with a sliding instability wasformed

(3) Using the mechanical model of the key stratumweighting interval the latter value was assessed at295m +is estimate complied with the physicalsimulation result of 30m Further the formula for

hydraulic support working resistance during theextraction of the working face with the considerationof cutting blocks was derived Calculations madeusing the engineering and geological parameters ofthe working face 42107 basically agreed with therespective field measurements

(4) +e basic countermeasures against strong groundpressure manifestations at the working face shouldweaken thick and hard rock roofs For this purposedirectional boreholes were drilled in the key stratumalong the two roadways and segmental hydraulicfracturing of the 22m-thick siltstone key stratumwasapplied Both the cyclic weighting intensity andinterval at the working face sharply dropped after thehydraulic fracturing +e absence of rib spalling orlarge roadway deformation proved that the direc-tional hydraulic fracturing could effectively weakenthe thick and hard roofs thus reducing the intensityof ground pressure manifestation at the workingface

Data Availability

+e data used for conducting classifications are availablefrom the corresponding author upon request

Conflicts of Interest

+e authors declare that there are no conflicts of interestregarding the publication of this paper

Acknowledgments

+e authors acknowledge the financial support for this workprovided by the National Natural Science Foundation ofChina (Grant no 51634007) and the Graduate InnovationFund Project of Anhui University of Science and Technologyof China (2019CX1003)

References

[1] T-b Zhao W-y Guo Y-l Tan C-p Lu and C-w WangldquoCase histories of rock bursts under complicated geologicalconditionsrdquo Bulletin of Engineering Geology and the Envi-ronment vol 77 no 4 pp 1529ndash1545 2017

303234363840424446

Wor

king

resis

tanc

e (M

Pa)

760 770 780 790 800 810 820 830 840750Advancing distance (m)

Post

The digitalpressure gauge

The infrared collector

Figure 11 Cyclic weighting interval of the working face after the fracturing of the key stratum

Shock and Vibration 11

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration

[2] T-b Zhao W-y Guo Y-l Tan Y-c Yin L-s Cai andJ-f Pan ldquoCase studies of rock bursts under complicatedgeological conditions during multi-seam mining at a depth of800 mrdquo Rock Mechanics and Rock Engineering vol 51 no 5pp 1539ndash1564 2018

[3] S L Yang J C Wang and J H Yang ldquoPhysical analogsimulation analysis and its mechanical explanation on dy-namic load impactrdquo Journal of China Coal Society vol 42no 2 pp 335ndash343 2017

[4] T Zhang Y X Zhao G P Zhu S L Wang and Z H Jiao ldquoAmulti-coupling analysis of mining-induced pressure charac-teristics of shallow-depth coal face in Shandong mining areardquoJournal of China Coal Society vol 41 no S2 pp 287ndash2962016

[5] H J Jiang S G Cao Y Zhang and C Wang ldquoStudy on thefirst failure and cavingmechanism of key strata of shallow coalseamrdquo Journal of Mining and Safety Engineering vol 33 no 5pp 860ndash866 2016

[6] J F Ju J L Xu andQ XWang ldquoCantilever structure movingtype of key strata and its influence on ground pressure in largemining height workfacerdquo Journal of China Coal Societyvol 36 no 12 pp 2115ndash2120 2011

[7] J F Ju J L Xu and W B Zhu ldquoInfluence of key stratacantilever structure motion on end-face fall in fully-mecha-nized face with super great mining heightrdquo Journal of ChinaCoal Society vol 39 no 7 pp 1197ndash1204 2014

[8] A M Suchowerska R S Merifield and J P Carter ldquoVerticalstress changes in multi-seam mining under supercriticallongwall panelsrdquo International Journal of Rock Mechanics andMining Sciences vol 61 pp 306ndash320 2013

[9] Q Qu J Xu R Wu W Qin and G Hu ldquo+ree-zonecharacterisation of coupled strata and gas behaviour in multi-seam miningrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 91ndash98 2015

[10] H Guo L Yuan B Shen Q Qu and J Xue ldquoMining-inducedstrata stress changes fractures and gas flow dynamics inmulti-seam longwall miningrdquo International Journal of RockMechanics and Mining Sciences vol 54 pp 129ndash139 2012

[11] G Si S Jamnikar J Lazar et al ldquoMonitoring and modellingof gas dynamics in multi-level longwall top coal caving ofultra-thick coal seams Part I borehole measurements and aconceptual model for gas emission zonesrdquo InternationalJournal of Coal Geology vol 144-145 pp 98ndash110 2015

[12] W Zhang D Zhang D Qi W Hu Z He and W ZhangldquoFloor failure depth of upper coal seam during close coalseams mining and its novel detection methodrdquo Energy Ex-ploration amp Exploitation vol 36 no 5 pp 1265ndash1278 2018

[13] Y Jiang H Wang S Xue Y Zhao J Zhu and X PangldquoAssessment and mitigation of coal bump risk during ex-traction of an island longwall panelrdquo International Journal ofCoal Geology vol 95 pp 20ndash33 2012

[14] Y L Tan X S Liu J G Ning and Y W Lu ldquoIn situ in-vestigations on failure evolution of overlying strata induced bymining multiple coal seamsrdquo Geotechnical Testing Journalvol 40 no 2 pp 1ndash14 2017

[15] P K Kaiser andM Cai ldquoDesign of rock support system underrockburst conditionrdquo Journal of Rock Mechanics and Geo-technical Engineering vol 4 no 3 pp 215ndash227 2012

[16] Z L Ge S H Li Z Zhou Y Y Lu B W Xia and J R TangldquoModeling and experiment on permeability of coal with hy-draulic fracturing by stimulated reservoir volumerdquo Rock Me-chanics and Rock Engineering vol 52 no 8 pp 2605ndash2615 2019

[17] J Liu C Liu and X Li ldquoDetermination of fracture location ofdouble-sided directional fracturing pressure relief for hard

roof of large upper goaf-side coal pillarsrdquo Energy Explorationamp Exploitation vol 38 no 1 pp 111ndash136 2020

[18] S F Yin H F Ma Z H Cheng et al ldquoPermeability en-hancement mechanism of sand-carrying hydraulic fracturingin deep mining a case study of uncovering coal in cross-cutrdquoEnergy Science amp Engineering vol 7 no 5 pp 1867ndash18812019

[19] Z Feng W Guo F Xu D Yang and W Yang ldquoControltechnology of surface movement scope with directional hy-draulic fracturing technology in longwall mining a casestudyrdquo Energies vol 12 no 18 p 3480 2019

[20] Z L Ge L Zhang J Z Sun and J H Hu ldquoFully coupledmulti-scale model for gas extraction from coal seam stimu-lated by directional hydraulic fracturingrdquo Applied Sciencesvol 9 no 21 p 4720 2019

[21] P Tan Y Jin B Hou L Yuan and Z Xiong ldquoExperimentalinvestigation of hydraulic fracturing for multi-type uncon-ventional gas co-exploitation in Ordos basinrdquo ArabianJournal for Science and Engineering vol 44 no 12pp 10503ndash10511 2019

[22] H K Han J L Xu X Z Wang J L Xie and Y T XingldquoMethod to calculate working surface abutment pressurebased on key strata theoryrdquo Advances in Civil Engineeringvol 2019 Article ID 7678327 20 pages 2019

[23] J Xie and J Xu ldquo+e corresponding relationship between thechange of goaf pressure and the key stratum breakingrdquoJournal of Geophysics and Engineering vol 16 no 5pp 913ndash925 2019

[24] X P Lai P F Shan J T Cao F Cui and H Sun ldquoSimulationof asymmetric destabilization of mine-void rock masses usinga large 3D physical modelrdquo Rock Mechanics and Rock En-gineering vol 49 no 2 pp 487ndash502 2016

[25] H C Li Similar Simulation Test of Mine Pressure ChinaUniversity of Mining and Technology Press Xuzhou China1988

[26] G X Xie J C Chang and K Yang ldquoInvestigations into stressshell characteristics of surrounding rock in fully mechanizedtop-coal caving facerdquo International Journal of Rock Mechanicsand Mining Sciences vol 46 no 1 pp 172ndash181 2009

[27] X Zhang P Gong K Wang J Li and Y Jiang ldquoCharac-teristic and mechanism of roof fracture ahead of the face in anLTCC panel when passing an abandoned roadway a casestudy from the Shenghua Coal Mine Chinardquo Rock Mechanicsand Rock Engineering vol 52 no 8 pp 2775ndash2788 2019

[28] XW Yin ldquoCutting block structure model of overburden withshallow buried coal seam and ultra-large mining heightworking facerdquo Journal of China Coal Society vol 44 no 7pp 1961ndash1970 2019

12 Shock and Vibration