An Innovative Support Structure for Gob-Side Entry ...

minerals Article

An Innovative Support Structure for Gob-Side Entry Retention in Steep Coal Seam Mining Jianguo Ning † , Jun Wang *,† , Tengteng Bu, Shanchao Hu * and Xuesheng Liu State Key Laboratory of Mining Disaster Prevention and Control Co-founded by Shandong Province and the Ministry of Science and Technology, Shandong University of Science and Technology, Qingdao 266590, China; [email protected] (J.N.); [email protected] (T.B.); [email protected] (X.L.) * Correspondence: [email protected] (J.W.); [email protected] (S.H.); Tel.: +86-532-8605-7330 (J.W. & S.H.) † These authors contributed equally to this work. Academic Editor: W. Scott Dunbar Received: 30 March 2017; Accepted: 8 May 2017; Published: 11 May 2017

Abstract: This study considered longwall working face No. 41101—located in a steeply inclined coal seam at the Awuzisu coal mine in Xinjiang, China—as an example in which macroscopic shear cracks had occurred in the cement-based filling body of the gob-side entry retention structure. A mechanical model of the support structure for the gob-side entry retention was first established. Then, field observations and laboratory tests were used to obtain the force exerted by the coal wall on the main roof, the relationship between the axial bearing capacity and compression ratio of the rubble inside the gob, the supporting force exerted by the rubble and filling body, and the thrust of the rubble on the filling body. The shear stress experienced by the roadside filling body of the gob-side entry retention in working face No. 41101 was calculated to be 15.89 MPa. To meet the needs of roadside support, an innovative roadside backfill–truss support structure was adopted, with a 60◦ angle of inclination used for the anchor bolts of the gob-side entry retention structure. In this way, the ultimate shear strength was improved by 107.54% in comparison with the cement-based filling body. Keywords: steeply inclined coal seam; gob-side entry retention; support mechanism; backfill–truss support

1. Introduction Over the past few decades, the gob-side entry retention technique has been widely used in China, Poland, and Ukraine [1–5]. A gob-side entry retention structure is a type of roadway without pillars, in which the former entry roadway is retained for transportation (or as a return air gate road) during the mining of the next panel by constructing an artificial wall along the gob-side, lagging behind the coal face being excavated [6,7]. As a major technique for realizing pillar-less mining in coal mines, it can improve the recovery rate of coal resources and prolong the mine’s service life. It presents a superior method for preventing major disasters in mines such as rock bursts and gas outbursts [8]. At present, gob-side entry retention has been successfully applied to the working faces of approximately horizontal, and gently inclined, coal seams [9–13]. However, it is still difficult to use gob-side entry retention for the working faces of steeply inclined coal seams because of complications. For example, in a steep coal seam, the rock beams of the lateral main roof (or the immediate roof strata) are likely to move along the inclined direction of the bedding [14–16]. Under such circumstances, the roadside filling body (an artificial wall created by pouring high-water quick-setting materials) is expected to bear a large shear force, which can lead to the overall slipping and bulging of the filling body [17–19]. Therefore, establishing a mechanical model for the support of the gob-side entry retention in steeply inclined coal seams and improving the anti-shear characteristics of the roadside Minerals 2017, 7, 75; doi:10.3390/min7050075

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filling body are crucial steps. This can not only provide reassurance when realizing gob-side entry Minerals 2017, 7, 75 2 of 18 retention in steep coal seams, but will also enrich and develop the theory and technique of gob-side entry retention. of the gob-side entry retention in steeply inclined coal seams and improving the anti-shear Acharacteristics mechanical model for thefilling support the gob-side retention steepreassurance coal seams was of the roadside bodyofare crucial steps. entry This can not only in provide when based realizing entry retention in steepconditions coal seams, of butthe will also mechanized enrich and develop the in the established ongob-side the geological and mining fully coalface theory and technique of gob-side entry retention. steeply inclined coal seam at the Awuzisu coal mine in Xinjiang, China. Then, this model was used to mechanical model for the support of the gob-side entry retention in steep coal seams was obtain the A shear stresses on the roadside filling body after the motion of the roof had stabilized, established based on the geological and mining conditions of the fully mechanized coalface in the and a suitable roadside backfill–truss support structure was proposed. Second, by performing steeply inclined coal seam at the Awuzisu coal mine in Xinjiang, China. Then, this model was used laboratory teststhe onshear similar materials, we acquired optimal installation angle anchor bolts to obtain stresses on the roadside fillingthe body after the motion of the roof for hadthe stabilized, used inand thearoadside backfill–truss support structure. Afterwards, the shear strength of the roadside suitable roadside backfill–truss support structure was proposed. Second, by performing backfill–truss support was calculated throughthe field testing. Finally,angle the system’s suitability laboratory tests onbody similar materials, we acquired optimal installation for the anchor bolts was in the roadside backfill–truss support structure. Afterwards, the shear strength of the roadside verifiedused on site. backfill–truss support body was calculated through field testing. Finally, the system’s suitability was verified on site. 2. Geological and Mining Conditions at Awuzisu Coal Mine in Xinjiang, China

The Awuzisu and coalMining mine Conditions is in the southeastern margin the Junggar 2. Geological at Awuzisu Coal Mine inofXinjiang, China Basin. Inclined shaft (including the main inclined shafts, auxiliary inclined shafts, and air shafts) development was adopted The Awuzisu coal mine is in the southeastern margin of the Junggar Basin. Inclined shaft in the coal field, as in Figure 1a,b.auxiliary The mineable coal seams #4, #5, and #6, from (including theshown main inclined shafts, inclined shafts, andinclude air shafts) development was which coal seam #4 isincurrently being With an 1a,b. average burial depth of 400include m, the#4, occurrence adopted the coal field, asmined. shown in Figure The mineable coal seams #5, and #6,of coal from which coal seam #4 isthe currently beingofmined. With an average seam #4 is stable. In addition, thickness the coal seam ranges burial from depth 1.8 m of to 400 2.2 m, m, the with the ◦ ◦ occurrence of coal seam #4 is stable. In addition, the thickness of the coal seam ranges from 1.8 m to angle average thickness being 2 m. The dip angle of the coal seam is 43 to 47 , with the average 2.2 m, with the average thickness being 2 m. The dip angle of the coal seam is 43° to 47°, with the ◦ of dip being 45 . The immediate roof is composed of mudstone, and is 3.8 m thick, with a uniaxial average angle of dip being 45°. The immediate roof is composed of mudstone, and is 3.8 m thick, compressive strength (USC) and density of 23 MPa and 2435 kg/m3 , respectively. The 24.6 m thick with a uniaxial compressive strength (USC) and density of 23 MPa and 2435 kg/m3, respectively. The 3 and USC of main roof is thick composed of is siltstone and fine sandstone. It has a density of 2784 kg/m 3 and 24.6 m main roof composed of siltstone and fine sandstone. It has a density of 2784 kg/m 55 MPaUSC (onofaverage). Figure 1c shows the lithological profile of the strata ininthe 55 MPa (on average). Figure 1c shows the lithological profile of roof the roof strata thecoal coal seam. seam.

+5 + 50 65 0 +4 +6 + 50 00 70 +4 00

0 +5 00

5 +3 0 00 +3

8 7 .3 1 0 2

50 +2 (a)

Figure 1. Cont.


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41101return airway

1500m

Shield support

150m

Face stop-line 41101 longwall face

Filling body

a

50m 100m Station I Station II 300m

a

( As shown in Table 1, the paste filling materials was used to construct roadway)

a Station III

50m 100m Station IV 280m

Advancing Support

41101 comprehensive transportation roadway

Rise head entire

Rise transportation entire

(The roadside backfill-truss support was used to construct roadway )

41103 ventilation roadway Main tailgate entire

(b)

(c)

(d)

Figure 1. Layout of working face No. 41101: (a) layout of working face; (b) layout of observation

Figurestations; 1. Layout of working face No. 41101: (a) layout of working face; (b) layout of observation (c) lithological profile of roof strata; and (d) section a-a. stations; (c) lithological profile of roof strata; and (d) section a-a. As the first working face in the first mining area of coal seam #4, working face No. 41101 uses fully coal face mining, withfirst strike and inclination of 1500 and 150 m, respectively. As themechanized first working in the mining area of lengths coal seam #4,mworking face No. 41101 uses Working face No. 41101 is supported by a ZZ6500/16/25 standing shield hydraulic support with a fully mechanized coal mining, with strike and inclination lengths of 1500 m and 150 m, respectively. rated working resistance of 6500 kN and setting load of 5232 kN. In addition, a MG450/1020-GWD Working face No. 41101 is supported a ZZ6500/16/25 electrical haulage shearer running onby variable frequency ACstanding is used toshield cut thehydraulic coal. With support a rated with a ratedvoltage working resistance of 6500 kN and setting load of 5232 kN. In addition, a MG450/1020-GWD of 3300 V and frequency of 50 Hz, the shearer can work at a web depth of 0.8 m. The shearer electrical shearer on less variable frequency AC is used coal. With has haulage an installed power running rating of not than 1006 kW, a haulage speed to of cut overthe 4.5 m/min, and aa rated with V a diameter of 2.5 m.of The is mined an average advancement pershearer voltageroller of 3300 and frequency 50working Hz, theface shearer canatwork at a web depth ofrate 0.8ofm.8 m The day. has an installed power rating of not less than 1006 kW, a haulage speed of over 4.5 m/min, and a roller The main transportation roadway of working face No. 41101 is excavated along the roof in an with a diameter of 2.5 m. The working face is mined at an average advancement rate of 8 m per day. irregular section, as illustrated in Figure 1d. The roadway is jointly supported by anchor bolts The main transportation roadway of working face No. 41101 is excavated along the roof in (anchor cables), a steel strip, and metal nets. Threaded steel resin bolts with a diameter and length of an irregular section, as illustrated inand Figure 1d. The roadway isreinforcement, jointly supported by anchor 20 mm and 1800 mm, respectively, containing no longitudinal are applied to the bolts (anchor cables), a steel strip, and metal nets. Threaded steel resin bolts with a diameter roof. The bolts, each of which is fixed using the CK2335 resin anchoring agent, are arranged and with length row and column lengths of 800 mm × 800 mm. The anchor cables, which are 15.2 mm in diameter of 20 mm and 1800 mm, respectively, and containing no longitudinal reinforcement, are applied to

the roof. The bolts, each of which is fixed using the CK2335 resin anchoring agent, are arranged with row and column lengths of 800 mm × 800 mm. The anchor cables, which are 15.2 mm in diameter and 7000 mm in length, are set at intervals of 1600 mm in both the rows and columns. In addition, three CK2335 resin anchoring agents are installed in each hole for the purpose of anchoring the ends. Threaded steel resin bolts without longitudinal reinforcement are used on the walls of the roadway.


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and 7000 mm in length, are set at intervals of 1600 mm in both the rows and columns. In addition, three CK2335 resin anchoring agents are installed in each hole for the purpose of anchoring the ends. Threaded bolts without longitudinal reinforcement are respectively, used on the walls of theatroadway. The bolts, steel withresin a diameter and length of 20 mm and 1800 mm, are fixed 800 mm The bolts,inwith diameter andcolumns. length ofIn20addition, mm andthe 1800 mm, respectively, are fixed atsteel 800with mm intervals bothathe rows and metal net is made of reinforced intervals in both the rows and columns. In addition, the metal net is made of reinforced steel with a diameter of 6.5 mm, and the size of the grid is 100 mm × 100 mm. The length, width, and thicknessa diameter 6.5 mm, size the and grid 5ismm, 100 mm × 100 mm. The length, width, and thickness of of the steelofstrips are and 1000the mm, 80ofmm, respectively. the steel strips areset 1000 mm, mm, of and mm,with respectively. Frames were into the 80 margin the5 gob, the space being filled with paste filling materials Frames set into the to margin of the the No. gob,41101 withmain the space being filled with as paste filling at every 3.2 mwere of mining length preserve transportation roadway the return materials 3.2 mface. of mining length to preserve No. 41101 transportation roadway as airway of at theevery working The cement-based filling the materials weremain prepared with the proportions the return airway of the working face. The cement-based filling materials were prepared with the listed in Table 1. The relationships between the strength and curing time of the cement-based filling proportions listed Table The relationships between the and strength curingas time the body are shown in in Figure 2. 1.Initially, the widths of the upper lowerand surfaces, well of as the cement-based body arewere shown in Figure Initially, theand widths the upper In and middle part of filling the filling body, designed to be2.1.5 m, 3.5 m, 2.5 m,of respectively. thelower field surfaces, as well as thebody middle themfilling body, were designed to be 1.5 m, 3.5 m, and 2.5 m, investigation, a filling thatpart wasof 300 long was prepared. respectively. In the field investigation, a filling body that was 300 m long was prepared. Table 1. Components and proportions of paste materials. Table 1. Components and proportions of paste materials. Stone Stone Granularity Granularity Early Strength Water-Reducing Early Strength Water-Reducing (mm) Water Cement Fly Ash Gangue River-Sand Stone Water Cement Fly Ash Gangue River-Sand Stone (mm) Agent Agent Agent Agent 227 340340 2.82.8 3.2 3.2 5–10 5–10 134 134 164 164 44 44 8585 227 Proportions of the Paste Filling Materials (kg/t) Proportions of the Paste Filling Materials (kg/t)

30

Shear strength of filling body(MPa)

Uniaxial compressive strength (Mpa)

35 30 25 20 15 10 5 0

0

5

10

15 20 Age (day)

(a)

25

30

35

25

20

15

10

5

0

0

2

4

6

8

10 12 14 16 18 20 22 24 26 28 30 32 34

Age(day)

(b)

Figure 2. Relationships between strengths and curing time of cement-based filling body: (a) Figure 2. Relationships between strengths and curing time of cement-based filling body: (a) relationship relationship between uniaxial compressive strength and age and (b) relationship between shear between uniaxial compressive strength and age and (b) relationship between shear strength and age. strength and age.

Two II)II) were established in the roadway, as shown in Figure 1b. A1b. laser Twoobservation observationstations stations(I(Iand and were established in the roadway, as shown in Figure A measuring instrument was used to measure the convergence between the roof and the floor, as well laser measuring instrument was used to measure the convergence between the roof and the floor, as as that the middle of the two of the roadway using theusing crossing shown in wellbetween as that between theparts middle parts of walls the two walls of the roadway themethod, crossingas method, as Figure 1d. In Figure 3, the results indicate that the roadway roof significantly sagged when the filling shown in Figure 1d. In Figure 3, the results indicate that the roadway roof significantly sagged when body was 40–45 m behind working i.e., five to i.e., six days after constructed. Moreover, the filling body was 40–45 the m behind theface, working face, five to six being days after being constructed. macroscopic shearing cracks occurred onoccurred the roadside filling body, and thebody, roadway to be Moreover, macroscopic shearing cracks on the roadside filling and needed the roadway reinforced many times to maintain the stability of the surrounding rocks. needed to be reinforced many times to maintain the stability of the surrounding rocks.


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Roof-to-floor at the station I Rib-to-rib at the station I Roof-to-floor at the station II Rib-to-rib at the station II

Convergence(mm)

200

Shear failure of filling boby

150

100

50

0

0

10

20

30

40

50

60

Distance behind 41101 longwall face (m) Figure 3. Deformation of surrounding rocks of gob-side entry retention work (cemented materials

Figure 3. Deformation of surrounding rocks of gob-side entry retention work (cemented materials were used to fill space in frameworks). were used to fill space in frameworks).

3. Shear Strength of Roadside Filling Body of Gob-Side Entry Retention Structure in Steep Coal

3. Shear SeamStrength of Roadside Filling Body of Gob-Side Entry Retention Structure in Steep Coal Seam 3.1. Mechanical Model of Gob-Side Entry Retention Work in Steep Coal Seam

3.1. Mechanical Model of Gob-Side Entry Retention Work in Steep Coal Seam

After a steep coal seam is exploited, the motion of the lateral roof of the working face can be separated into two stages [20,21]. During stage I, the immediate roof strata subside andcan be After a steep coal seam is exploited, the motion of the lateral roofrotationally of the working face collapse into the gob. This collapsed rubble slips downward along the inclined direction of the coal separated into two stages [20,21]. During stage I, the immediate roof strata rotationally subside and seam into under thegob. effectThis of gravity, and gradually forms heaps at the side of roadsidedirection filling body collapse the collapsed rubble slips downward along thethe inclined of of the coal the gob-side entry retention work. In stage II, when the hanging length of the rock beam of the seam under the effect of gravity, and gradually forms heaps at the side of the roadside filling body of lateral main roof reaches its limit, the rock beam is fractured in the coal wall and rotationally the gob-side entry retention work. In stage II, when the hanging length of the rock beam of the lateral subsides until it contacts, and compacts, the rubble inside the gob, as shown in Figure 4. main roof the rock beam is fractured in the coal wall and rotationally subsides Inreaches contrastits to limit, approximately horizontal and gently inclined coal seams, the roof strata of theuntil it contacts, and compacts, the rubble inside the gob, as shown in Figure 4. working face of a steeply inclined coal seam tend to move downward along the inclined direction. In contrastpushed to approximately inclined coal seams, roof strata Meanwhile, by the rubblehorizontal inside the and gob, gently the roadside filling body of thethe gob-side entry of the working face bears of a steeply coalstress, seamwhich tend tomakes moveitdownward the inclined retention a largeinclined tangential easier for along the filling body to direction. be shear-fractured. Assume that the dip angle of the coal seam is α, and the rock beam A of the lateral Meanwhile, pushed by the rubble inside the gob, the roadside filling body of the gob-side entry main roof subsides aroundwhich point Dmakes of the itfracture andfilling is pushed bytorock beam B at retention bears a largerotationally tangential stress, easier line for the body be shear-fractured. point E. The rotation angle of rock beam A is θ, and there is no separation between the main Assume that the dip angle of the coal seam is α, and the rock beam A of the lateral mainroof roofand subsides the immediate roof. The self-weight forces of the immediate roof and rock beam A act on the rubble rotationally around point D of the fracture line and is pushed by rock beam B at point E. The rotation inside the gob, the filling body, and the coal wall. The mechanical models of the roof and filling body angle of rock beam A is θ, and there is no separation between the main roof and the immediate roof. of the gob-side entry retention in a steeply inclined coal seam can be established in this way, as The shown self-weight forces of5,the immediate roof and rock beam A act on the rubble inside the gob, in Figures 4 and respectively.

the filling body, and the coal wall. The mechanical models of the roof and filling body of the gob-side entry retention in a steeply inclined coal seam can be established in this way, as shown in Figures 4 and 5, respectively.


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EE FFgy

GGEE gy

FFcy FFc c cy

FFg g

GGz zFFbyby FFbb Roadway Roadway

αα AA

qq1y 1y qq1x 1x

qq2y 2y qq2x 2x

qqby BB by

L0

L0 AA

dd

BB

qqgy gy

CC qqgx gx

qqbx bx LL3 3

CC

Figure 4.4.4. Mechanical model ofofof roof ofofof gob-side entry retention work ininin steeply inclined coal seam. Figure Mechanical model roof gob-side entry retention work steeply inclined coal seam. Figure Mechanical model roof gob-side entry retention work steeply inclined coal seam.

′′ FFby by

′′ FFbx bx

Figure 5.5.Mechanical model ofoffilling body entry work steeply inclined Figure Mechanical model filling bodyof ofgob-side gob-side entryretention retention workininwork steeply inclinedcoal coalseam. seam. Figure 5. Mechanical model of filling body of gob-side entry retention in steeply inclined coal seam.

3.1.1. 3.1.1.Supporting SupportingForce Forceof ofCoal CoalWall Wall

3.1.1.IfIfSupporting of supporting Coal Wall force we that weassume assumeForce thatthe the supporting forceof ofthe thecoal coalwall wallon onthe theroof roofstrata strataisisFFc c(Figure (Figure4), 4),then thenits its component in the y-direction, F cy, is linearly distributed and varies from q1y to q2y, which can be component in thethat y-direction, Fcy, is force linearly distributed varies 1y F to q2y, which can its be If we assume the supporting of the coal walland on the roof from strataqis 4), then c (Figure obtained using aa borehole stress meter. Meanwhile, its component in the x-direction, FFcxcx, , isis also obtained using borehole stress meter. Meanwhile, its component in the x-direction, also component in the y-direction, Fcy , is linearly distributed and varies from q1y to q2y , which can be linearly and from and the from wall on roof linearlydistributed distributed andranges ranges fromqq1x1x andqq2x2x. .Then, Then, theforces forcesacting acting fromthe thecoal coal onthe the roof obtained using a borehole stress meter. Meanwhile, its component in the x-direction, Fcxwall , is also linearly strata in the xand y-directions, F cx and Fcy, respectively, can be calculated using Equation (1): strata in the xand y-directions, F cx and F cy , respectively, can be calculated using Equation (1): distributed and ranges from q1x and q2x . Then, the forces acting from the coal wall on the roof strata in 11 the x- and y-directions, Fcx and Fcy , respectively, can be calculated using Equation (1): FF = LL0 ( q( q1y ++qq2y ) ) cy = cy 2y 22 0 1y

Fcy = 12 L01(1q1y + q2y ) F = L（ q1x αα−−θθ) ) , cy / /tan( =FF tan( L0（ q1x+q +qq2x2x)))== 0 + Fcx = Fcx21cxL=02(2q1x Fcycy / tan (α, − θ ) 2x ,

(1) (1) (1)

; ;A αα− isisthe of the seam (°); θθisisthe where (Δ ==hh−= −−z1)1) where arcsin −KA m (K 1the ); αdip is angle the dip of the coal (◦ );rotation θ is the dip angle ofangle thecoal coal seam (°);seam the rotation where θθθ== =arcsin arcsin ( Δhh(/∆h/L /LL2 2) ); ; 2Δ)Δh; h∆h −mmzhz( (K A ◦ rotation anglebeam of rock beam Amain in the main roof ( the ); L2broken is the broken length ofbeam rock beam AΔ (m); ∆h is angle of rock A in the roof (°); L 2 is length of rock A (m); angle of rock beam A in the main roof (°); L2 is the broken length of rock beam A (m); Δhh isisthe the the deflection (subsidence) of rock beam A (m); h is the mining height of the working face (m); m z and z and deflection (subsidence) of rock beam A (m); h is the mining height of the working face (m); m deflection (subsidence) of rock beam A (m); h is the mining height of the working face (m); mz andLL0 0 L0 are the thicknesses of the immediate roof the and the length fromlateral the lateral fracture line to the coal are arethe thethicknesses thicknessesof ofthe theimmediate immediateroof roofand and thelength lengthfrom fromthe the lateralfracture fractureline lineto tothe thecoal coalwall, wall, wall, respectively (m); and K is the bulking coefficient of the gangue, and its value is taken from the respectively A respectively (m); (m);and and KKAA isis the the bulking bulking coefficient coefficient of of the the gangue, gangue, and andits its value value isis taken takenfrom from the the relevant literature literature [22]. [22]. relevant relevant literature [22]. 3.1.2. 3.1.2.Supporting SupportingForce Forcefrom fromRubble Rubbleinside insideGob Gob


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3.1.2. Supporting Force from Rubble inside Gob The components of the supporting force of the rubble, Fg , in the x- and y-directions are Fgx and Fgy , respectively, where the component in the x-direction, Fgy , is related to the contacting length of rock beam A with the rubble, and the final compression of the rubble. In a steeply inclined coal seam, because the rubble inside the gob mainly accumulates around the roadside filling body, rock beam A can make sufficient contact with it. The length L3 of the rubble in contact can be calculated using Equation (2): L3 = L2 − L1 , (2) where L1 is the strike length from the lateral fracture line to the end of the immediate roof in meters. Because rock beam A makes sufficient contact with the rubble inside the gob, the supporting force of the rubble on rock beam A can be assumed to be a uniform load qgy . In accordance with a related study [23], it can be found that qgy is exponentially related to the compression ratio of the rubble, as follows: . qgy = a exp(bε) + c, (3) .

where a, b, and c are constants obtained from an experiment, ε is the compression ratio of the rubble, . ε = S/KA MZ , S is the final compression of the rubble inside the gob, and S = (KA − 1) Mz . Therefore, the supporting forces from the rubble inside the gob to rock beam A, in the x- and y-directions Fgx and Fgy , respectively, are represented as follows: .

Fgy = qgy L3 = [ a exp(bε) + c] L3 . Fgx = qgx L3 = Fgy / tan(α − θ )

(4)

3.1.3. Supporting Force from Filling Body The equation for the mechanical equilibrium in the x- and y-directions is established according to the model shown in Figure 4. Based on this equation, the forces acting between the roadside filling body and roof strata, Fby , and Fbx , can be calculated as follows: Fby = L1 mz γz cos(α − θ ) + L2 mE γE cos(α − θ ) + FEy − FDy − Fcy − Fgy ,

(5)

Fbx = L1 mz γz sin(α − θ ) + L2 mE γE sin(α − θ ) + FEx − FDx − Fcx − Fgx ,

(6)

where γz and γE are the bulk densities of the immediate roof and main roof (kN/m3 ), respectively; mE is the thickness of the rock beam of the main roof (m); FDy and FDx are the forces caused by rock beam A at articulation points D; and FEx and FEy are the forces caused by rock beam B at articulation point E. According to the result of Tan et al. [2,7], FEy − FDy ≈ 0 and FEx − FDx ≈ 0. 3.1.4. Shear Stress on Roadside Filling Body 0 and F 0 represent the In the mechanical model of the roadside filling body shown in Figure 5, Fby bx forces acting between the roof strata and filling body in the y- and x-directions, respectively. They are the force and reaction force to Fby and Fbx , respectively. F0 refers to the supporting force from the floor to the filling body, while f 1 is the friction force between the filling body and floor strata. Suppose that the thrust caused by the rubble inside the gob on the filling body is linearly distributed, varies from q3 to q4 , and can be measured by the in situ stress sensor. By establishing the equation of mechanical equilibrium in the x-direction, we can obtain the shear stress τ on the roadside filling body as follows:

τ=

0 − 2f hc (q3 + q4 ) + 2G sin α+2Fbx 1 , 2d

(7)

Minerals2017, 2017,7,7,75 75 Minerals

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where f1 = μ1 F0 = μ1 ( Fby′ +0 G cos α ) , μ1 is the friction coefficient between the filling body and floor where f 1 = µ1 F0 = µ1 ( Fby + G cos α), µ1 is the friction coefficient between the filling body and floor strata, strata,GGisisthe theself-weight self-weightof ofthe thefilling filling body body (kN/m), (kN/m),ddisisthe theequivalent equivalentwidth widthofofthe theroadside roadsidefilling filling body (m), and h c is the vertical height of the filling body in the y-direction (m). body (m), and hc is the vertical height of the filling body in the y-direction (m). 3.2.Shear ShearStrength StrengthofofFilling FillingBody Body 3.2. 3.2.1.Supporting SupportingForce Forcefrom fromCoal CoalWall Wall 3.2.1. Two observation observationstations stationswere wereestablished establishedatat100 100m m(station (stationI)I)and and150 150m m(station (stationII) II)from fromthe the Two open-offcut cutin inthe the roadway, roadway, as as shown shown in in Figure Figure 6a. 6a. Four Fourborehole boreholestress stressmeters meterswere werefixed fixedin inthe the open-off coal wall of the observation stations at depths of 2 m, 4 m, 6 m, and 8 m, respectively. These borehole coal wall of the observation stations at depths of 2 m, 4 m, 6 m, and 8 m, respectively. These borehole stress meters meters were 2m to monitor the the changes in coal over aover 30 day period, stress were set setatatintervals intervalsofof 2m to monitor changes in pressure coal pressure a 30 day as shown in Figure 6b,c. The changes in thein pressure on theonlateral coal walls in theinroadway after period, as shown in Figure 6b,c. The changes the pressure the lateral coal walls the roadway the working face was werewere found by averaging the data usingusing the borehole stress after the working face advanced was advanced found by averaging the obtained data obtained the borehole meters at the same depth (Figure 7). According to another study [2,24], the fracture line of the stress meters at the same depth (Figure 7). According to another study [2,24,], the fracture line ofmain the roof was the at same as the peak abutment pressure.pressure. Therefore, the lateralthe peak abutment main roofatwas thelocation same location as the peak abutment Therefore, lateral peak pressure of the working was situated 4 m inside theinside coal wall of the roadway, abutment pressure of the face working face wasapproximately situated approximately 4m the coal wall of the namely, atnamely, L0 = 4 m. forsolving the average values of data acquired the two observation roadway, at LBy 0 =solving 4 m. By for the average values of datafrom acquired from the two stations, thestations, vertical pressures at the fractureatline the main roof the roadway surface, q1y and observation the vertical pressures theof fracture line of near the main roof near the roadway q , were calculated to be 0.5 MPa and 1.5 MPa, respectively. 2y surface, q1y and q2y, were calculated to be 0.5 MPa and 1.5 MPa, respectively. Because the the mining mining height height of of working working face face No. No. 41101 41101 isis 22 m m (on (on average), average), the the deflection deflection Because ◦ could finally be (subsidence) of rock beam A of the main roof was 1.126 m. Thus, a value of 3 (subsidence) of rock beam A of the main roof was 1.126 m. Thus, a value of 3° could finally be obtainedfor forthe therotation rotationangle angleθθ of of rock rock beam beam A, A, and andthe thedip dipangle angleof ofthe thecoal coalseam, seam,α,α,was was45°. 45◦ . obtained Accordingto to Equation Equation(1), (1), the the forces forces acting acting from from the the coal coal wall wall to to the the main main roof roof strata, strata, F Fcx and FFcycy, , cx and According 3 kN and 4.00 × 10 3 kN, respectively. 3 3 were 4.44 × 10 were 4.44 × 10 kN and 4.00 × 10 kN, respectively.

(a)

4m

d

Pressure sensors 1

2

8m

hc

6m

Stake 3

2m

2m

4

Borehole stressmeter 1-Observation line

2- telescoping rod

Coal seam D

3-Edge of the entry 4-Borehole stress

(b)

(c)

Figure 6. Layout of field monitoring stations: (a) layout of observation stations; (b) profile of Station Figure 6. Layout of field monitoring stations: (a) layout of observation stations; (b) profile of Station I; I; and (c) layout of borehole stress-meters. and (c) layout of borehole stress-meters.


9 of 18 Station Ⅰ Station Ⅱ

1.6 1.4


Stress(MPa)


9 of 18

9 of 18

1.2 1.0

Station Ⅰ Station Ⅱ

1.6 0.8 1.4

Stress(MPa)

0.6 1.2 0.4 1.0

2

0.8

4

6

8

Distance between broehole stressmeter(m)

Figure 7. Pressure curves in coal mass.

0.6 0.4

3.2.2. Supporting Force of Rubble inside Gob 2

1.

4

6

Distance between broehole stressmeter(m) experiment on load-carrying properties of rubble inside gob

8

Figure 7. in mass. Figure 7. Pressure Pressure curves in coal coal mass. To explore the load-carrying properties ofcurves the rubble inside the gob, the author designed loading molds, which were made of Q350 stainless steel plates measuring 300 mm × 300 mm × 300 3.2.2. of inside Gob 3.2.2. Supporting Force of Rubble Rubble inside Gob mm. Supporting The rubble,Force randomly collected on site, was placed into the molds, with fixed lateral constraints. Thus, the load-carrying properties of the inside obtained 1. experiment experiment on on load-carrying load-carrying properties properties of of rubble rubble gobrubble were similar to those under 1. inside gob actual conditions. Figure 8 shows the device and molds used in the loading test. The loading test was To explorethe theload-carrying load-carrying properties of rubble the rubble thethe gob, the designed author designed To explore properties of the insideinside the gob, author loading performed on a WE-1000C hydraulic testing machine under multi-stage loading. Meanwhile, by loading molds, which were made of Q350 stainless steel plates measuring 300 mm × 300 mm 300 molds, which made Q350 at stainless steel plates measuring 300 mm tape × 300 mm × 300×mm. measuring thewere height of theofrubble each load increment using a measuring with a precision of mm.rubble, The rubble, randomly collected on site, wasinto placed into the molds, with fixed lateral The randomly collected on site, was placed the molds, with fixed lateral constraints. 0.01 mm, the relationship between the bearing capacity and compression ratio of the gangue was constraints. Thus, the load-carrying properties of the obtained rubble were under similaractual to those under Thus, the load-carrying properties of results the obtained were similar to those conditions. acquired, as shown in Figure 9. The of thisrubble experiment suggested that the bearing capacity of actual conditions. Figure 8 shows the device and molds used in the loading test. The loading test was Figure 8 shows thegob device molds usedwith in the Thecompression loading test ratio was performed on the rubble in the grewand exponentially anloading increasetest. in the of the rubble. performed onhydraulic a WE-1000C hydraulic testing under multi-stage loading. by aAfter WE-1000C testing machinebetween undermachine multi-stage loading. Meanwhile, by Meanwhile, measuring fitting Equation (3), the relation the axial bearing capacity and compression ratiothe of measuring therubble height at of each the rubble at each loadusing increment using a tape measuring with a of precision of height of the load increment a measuring with atape precision 0.01 mm, the gangue can be expressed as follows: 0.01 mm, the relationship between thecapacity bearingand capacity and compression ratio of thewas gangue was the relationship between the bearing compression ratio of the gangue acquired, qgyresults =0.02468 exp(8.427 ε ) − 0.0074 acquired, as shown in Figure 9. The of this experiment suggested that the bearing capacity of (8) as shown in Figure 9. The results of this experiment suggested that the . bearing capacity of the rubble the rubble in theexponentially gob grew exponentially with in an the increase in the compression ratio ofAfter the rubble. in the gob grew with an increase compression ratio of the rubble. fitting The supporting force from the rubble in the theaxial gob bearing to rock capacity beam Aand of the main roofratio in the After fitting Equation (3), the relation between compression of Equation (3), the relation between the axial bearing capacity and compression ratio of the gangue can y-direction is asbe follows: the gangue can expressed as follows: be expressed as follows: .

F [0.02468 qgy 0.02468exp(8.427 exp(8.427εεε))−−−0.0074] 0.0074.L3 . qgy== =0.02468 exp(8.427 0.0074 gy .

(8) (9) (8)

The supporting force from the rubble in the gob to rock beam A of the main roof in the Mould y-direction is as follows: Fgy = [0.02468 exp(8.427ε ) − 0.0074]L3

.

Mould

Figure 8. 8. Device Device and and molds molds used used in in loading loading test test of of rubble. rubble. Figure

Figure 8. Device and molds used in loading test of rubble.

(9)


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0.7

Axial bearing capacity (MPa)

0.6 0.5 0.4 0.3 0.2 0.1 0.0

0

1

2

3

compression ration of gangues (%)

4

Figure 9. Curve relating bearing capacity to compression ratio of rubble in gob.

Figure 9. Curve relating bearing capacity to compression ratio of rubble in gob. 2.

supporting force from rubble in gob

The supporting force from the rubble in the gob to rock beam A of the main roof in the y-direction A laser measuring instrument was installed at the end of the support of the working face to is as follows: measure the fracture span of the immediate roof in. the gob. This field test revealed that the Fgy =the [0.02468 (8.427 ε)body. − 0.0074 ] L3 . span L2 of the rock beam (9) immediate roof collapsed along outside exp of the filling The fracture 2.

of the lateral main roof was approximately equal to the periodic weighted length of the working face supporting force from rubble in gob

[24,25], which was 21.5 m. The length from the lateral fracture line to the end of the immediate roof,

L1 was 8.5 m. According to Equation (2), the length of rock A support of the main was inface to A laser measuring instrument was installed at the endbeam of the of roof the that working contact with the rubble, L 3, was 13 m. In addition, the immediate roof was 3.8 m thick, and the measure the fracture span of the immediate roof in the gob. This field test revealed that the immediate bulking coefficient KA of the rubble inside the gob was 1.3. The final compression S of the rubble roof collapsed along the outside of the filling body. The fracture span L2 of the rock beam of the lateral inside the gob was 1.14, with a corresponding compression ratio of 0.19. Therefore, based on main roofEquations was approximately to theforces periodic of the [24,25], (4) and (9), theequal supporting fromweighted the rubblelength inside the gob working to the rockface beam A, Fgx which 3 kN/m, was 21.5and m. FThe length from the lateral to the of the immediate roof, L1 was 8.5 m. gy, can be calculated to be 1.65 × fracture 103 kN/m line and 1.49 × 10end respectively. According to Equation (2), the length of rock beam A of the main roof that was in contact with the Supporting Force from Filling Body rubble, L3.2.3. 3 , was 13 m. In addition, the immediate roof was 3.8 m thick, and the bulking coefficient Δh of KA of the rubble inside final S of the subsidence rubble inside therock gob was Working face the No. gob 41101was is 2 1.3. m in The height (on compression average). The deflection beam A of the main roof was 1.126 m. Based thebased rotationon angle θ of rock (4) beam 1.14, with a corresponding compression ratio on of these 0.19.parameters, Therefore, Equations and (9), A was found to be 3°. As previously mentioned, the bulk densities of the immediate and main roofs the supporting forces from the rubble inside the gob to the rock beam A, Fgx and Fgy , can be calculated kN/m3 and 27.28 kN /m33, respectively; the thickness mE of the rock beam of the main roof is to be 1.65are ×23.86 103 kN/m and 1.49 × 10 kN/m, respectively. 24.6 m, and the dip angle α of the coal seam is 45°. According to Equation (5), the vertical supporting forces Fby and Fbx from the roadside filling body on the roof strata are 5.76 × 103 kN/m and 4.09 × 103 3.2.3. Supporting Force from Filling Body kN/m, respectively.

Working face No. 41101 is 2 m in height (on average). The deflection subsidence ∆h of rock inside Gob on Filling Body beam A 3.2.4. of theThrust mainfrom roofRubble was 1.126 m. Based on these parameters, the rotation angle θ of rock beam ◦ After the filling body was constructed, pressure sensors were fixed distances of 500and mm main from roofs A was found to be 3 . As previously mentioned, the bulk densities of at the immediate 3 3 the roof and floor strata on one side of the filling body at observation stations I and II, as shown in main are 23.86 kN/m and 27.28 kN /m , respectively; the thickness mE of the rock beam of the Figure 6b. These sensors were numbered I-1, I-2, II-1, and II-2, where the Roman numerals show the ◦ roof is 24.6 m, and the dip angle α of the coal seam is 45 . According to Equation (5), the vertical observation stations and the Arabic numerals represent the serial numbers of the pressure sensors. supporting forces Fbx from roadside body on theand roof strataagainst are 5.76 103 kN/m by and Under the F effect of gravity, thethe rubble insidefilling the gob collapsed pushed the×stress 3 and 4.09 sensors. × 10 kN/m, respectively. In this way, we found the thrust exerted by the rubble inside the gob on the filling body (Figure 10).

3.2.4. Thrust from Rubble inside Gob on Filling Body After the filling body was constructed, pressure sensors were fixed at distances of 500 mm from the roof and floor strata on one side of the filling body at observation stations I and II, as shown in Figure 6b. These sensors were numbered I-1, I-2, II-1, and II-2, where the Roman numerals show the observation stations and the Arabic numerals represent the serial numbers of the pressure sensors.


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Under the effect of gravity, the rubble inside the gob collapsed and pushed against the stress sensors. In this way,2017, we 7,found the thrust exerted by the rubble inside the gob on the filling body (Figure Minerals 75 11 of 18 10). 1.6

I-1 sensor I-2 sensor II-1 sensor II-2 sensor

Stress (MPa)

1.4 1.2 1.0 0.8 0.6 0.4 0.2

0

10

20

30

40

50

60

70

Distance behind 41101 longwall face (m)

80

90

Figure 10. Thrust from rubble inside gob on filling body.

Figure 10. Thrust from rubble inside gob on filling body.

As can be seen from Figure 10, with the continuous advance of the working face, the thrust from the rubble the gob on the10, filling slowly increased. Moreover, values recorded As can beinside seen from Figure withbody the continuous advance of thethe working face, theatthrust sensors #1 and #2 stabilized at 1.5 MPa and 1.1 MPa, respectively, at 45 m behind the working face. from the rubble inside the gob on the filling body slowly increased. Moreover, the values recorded at Thus, it could be found that the thrusts from the rubble inside the gob on the filling body, q 3 and q4, sensors #1 and #2 stabilized at 1.5 MPa and 1.1 MPa, respectively, at 45 m behind the working face. were 1.1 MPa and 1.5 MPa (on average), respectively.

Thus, it could be found that the thrusts from the rubble inside the gob on the filling body, q3 and q4 , were3.2.5. 1.1 MPa 1.5 MPa (on average), Shearand Strength of Roadside Fillingrespectively. Body According toof theRoadside literature Filling [26], theBody frictional angle between the filling body and floor strata in 3.2.5. Shear Strength the roadway was 40°, and the vertical height hc in the y-direction, equivalent width d, and bulk According to the literature [26],2 the frictional the filling body strata in the density of the filling body were m, 2.5 m, andangle 15.25 between kN/m3, respectively. Basedand on floor the obtained ◦ roadway the vertical height hresults, y-direction, equivalent width d,filling and bulk of field was data 40 and, and laboratory experimental shear stress τ on the roadside bodydensity was c in thethe calculated be 15.89 Equation (7). 3 , respectively. Based on the obtained field data and the filling bodytowere 2 m,MPa 2.5 using m, and 15.25 kN/m

laboratory experimental results, the shear stress τ on the roadside filling body was calculated to be Discussion 15.894.MPa using Equation (7). Compared with that of a gently inclined coal seam, the roof strata of the working face of the

4. Discussion steeply inclined coal seam tended to move in the inclined direction. When performing gob-side entry retention in the steeply inclined coal seam, the roadside filling body was mainly subjected to Compared with that of a gently inclined coal seam, the roof strata of the working face of the steeply shear forces. When the lateral immediate roof collapsed to become rubble inside the gob and the inclined coal seam tended to move in the inclined direction. When performing gob-side entry retention in main roof fractured, the filling body beside the remaining roadway was subjected to a low shear the steeply inclined seam, the roadside filling bodyand wascompact mainly the subjected forces. stress. When thecoal main roof began to subside, contact, rubble,to theshear rubble insideWhen the the lateral immediate roof collapsed to become rubble inside the gob and the main roof fractured, the gob pushed the filling body, under pressure from the roof. During this stage, the filling body wasfilling bodybearing beside athe remaining roadway was subjected to athe low shear stress. When thelarger mainthan roofitsbegan large shear stress. If the shear stress τ from roadside filling body was to subside, and compact thefailure rubble,would the rubble the gob pushed the fillingwith body, ultimatecontact, shear strength τU, shear occur inside in the filling body. In accordance theunder thrust from theroof. rubble insidethis No.stage, 41101’s producing a working on the filling body (seeshear pressure from the During thegob filling body was bearingforce a large shear stress. If the the deformation of the rocks in the shear remaining roadway alongfailure the gob stressFigure τ from10) theand roadside filling body wassurrounding larger than its ultimate strength τU , shear would (see Figure 3), the main roof began to rotate, subside, and compact the rubble inside the gob at occur in the filling body. In accordance with the thrust from the rubble inside No. 41101’s gob 40–45 producing m behind the working face. By combining the mechanical models, it was found that the roadside a working force on the filling body (see Figure 10) and the deformation of the surrounding rocks in the filling body was subjected to a shear stress of 15.89 MPa. No. 41101’s working force was advanced by remaining roadway along the gob (see Figure 3), the main roof began to rotate, subside, and compact 7–8 m per day (on average). When the filling body was 40–45 m behind the working force, it had the rubble inside the gob at 40–45 behind the working face. By combining the body mechanical models, been constructed for five or six m days. The maximum shear strength of the filling was 12.38 it was found that the roadside filling body was subjected to a shear stress of 15.89 MPa. No. 41101’s MPa, as demonstrated in Figure 2b, which could not satisfy the field requirement for the shear working force Therefore, was advanced by 7–8 m per day (ongenerally average).occurred When the bodybody was at 40–45 mm behind strength. macroscopic shear cracks on filling the filling 40–45 the working force, it had face, beenand constructed for five six days. The maximum shear the filling behind the working the roadway wasorsignificantly deformed. Thus, thestrength roadsideoffilling in the MPa, steeply coal seam needed2b, to provide a largenot supporting force the roof strata bodybody was 12.38 asinclined demonstrated in Figure which could satisfy the fieldtorequirement for the and have a high shear strength to adapt to the roof movement and thrust caused by the shear strength. Therefore, macroscopic shear cracks generally occurred on the filling rubble. body at 40–45 m

behind the working face, and the roadway was significantly deformed. Thus, the roadside filling body


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in the steeply inclined coal seam needed to provide a large supporting force to the roof strata and have Minerals 2017, 7, 75 12 of 18 a high shear strength to adapt to the roof movement and thrust caused by the rubble. To improve the shear-resistance of the filling body, we proposed a roadside filling body–truss To improve the shear-resistance of the filling body, we proposed a roadside filling body–truss support structure, as shown inin Figure supportstructure structure is composed of three anchor support structure, as shown Figure11. 11.This This truss truss support is composed of three anchor bolts bolts and and a steel strip. Therein, two inclined anchor bolts (B1 and B2) are anchored between the a steel strip. Therein, two inclined anchor bolts (B1 and B2) are anchored between the roof roof and floor, whilewhile the third anchor boltbolt (B3)(B3) is fixed in the filling body and lies parallel and floor, the third anchor is fixed in the filling body and lies paralleltotothe thecoal coalseam. three anchor bolts are fastened to the steel strip through the connector. This support Theseseam. threeThese anchor bolts are fastened to the steel strip through the connector. This support structure structurelateral can provide lateral support resistances 1, and for thebody, filling among body, among can provide support resistances qL, F1 , andqL, F2Ffor theF2filling whichwhich qL is qL theis force the force acting on the steel strip on the roadside filling body in the form of a uniform load, F1 acting on the steel strip on the roadside filling body in the form of a uniform load, F1 represents represents the constraining force provided by the inclined bolts B1 and B2 to the filling body, and F2 the constraining force provided by the inclined bolts B1 and B2 to the filling body, and F2 is the is the constraining force offered by bolt B3 to the filling body. Meanwhile, the truss support structure constraining force offered by bolt B3 to the filling body. Meanwhile, the truss support structure anchors anchors the filling body with the roof and floor strata, which improves the shear strength of the the filling body with the roof and floor strata, which improves the shear strength of the filling body. filling body.

Figure 11. Filling body–truss support structure beside remaining roadway along gob.

Figure 11. Filling body–truss support structure beside remaining roadway along gob.

5. Laboratory Experiment on Shear Performance of Backfill–Truss Support Structure

5. Laboratory Experiment on Shear Performance of Backfill–Truss Support Structure 5.1. Experimental Materials

5.1. Experimental Materials

Based on the similarity principle, steel wires and a thin piece of sheet steel (45 steel) were used to simulate thesimilarity anchor bolts and steel steel strip, wires respectively. property of the used Based on the principle, and a The thinmechanical piece of sheet steelparameters (45 steel) were steel wires and thin piece of sheet steel are listed in Table 2. A mixture of barite powder and rosinof the to simulate the anchor bolts and steel strip, respectively. The mechanical property parameters alcohol solution was used as the anchoring agent. In addition, the support structure had a geometric steel wires and thin piece of sheet steel are listed in Table 2. A mixture of barite powder and rosin similarity of was 1:10.used Tableas3the listsanchoring the diameters of In theaddition, anchor bolts and boreholes, along the alcohol solution agent. the support structure hadwith a geometric intervals and lengths of the anchor bolts, while Table 4 lists the geometrical parameters of the steel similarity of 1:10. Table 3 lists the diameters of the anchor bolts and boreholes, along with the intervals strip.

and lengths of the anchor bolts, while Table 4 lists the geometrical parameters of the steel strip. Table 2. Mechanical parameters of anchor bolts, steel strip, and 45 steel.

Table 2. Mechanical parameters of anchor bolts, steel strip, and 45 steel.

Parameters Element

Parameters Rock bolt Element

SteelRock stripbolt Steel wires, thinstrip sheet Steel steel Steel wires, thin

sheet steel

Tension (MPa)

Shear Strength (MPa)

Anchoring Force (MPa)

Elastic Modulus (GPa)

Extensibility

Tension 200–600 (MPa) 350–400 200–600 350–400 600

Shear Strength 260–600 (MPa) / 260–600 / 400

Anchoring ≥50 Force (MPa) ≥50/ / 80

Elastic 200 Modulus (GPa) 200 200 200 210

≥16% ≥16% ≥16% ≥16%

600

400

80

210

≥16%

Extensibility ≥16%

Table 3. Geometrical parameters of truss support structure. Parameters Element Engineering prototype Truss support structure

Borehole Diameter (mm) 30 3

Bolt Diameter (mm) 20 2

Space (mm) 800 80

Bolt Length (mm) Bolt B1 and Bolt B2 Bolt B3 3000 1800 300 180


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Table 3. Geometrical parameters of truss support structure. Parameters Element Engineering prototype Truss support structure Minerals 2017, 7, 75

Borehole Diameter (mm)

Bolt Diameter (mm)

Space (mm)

30 3

20 2

800 80

Bolt Length (mm) Bolt B1 and Bolt B2 3000 300

Bolt B3 1800 180 13 of 18

Table 4. Geometrical parameters of steel strip.

Type

Width (mm) Thickness (mm) Length (mm) Space (mm) Table 4. Geometrical parameters of steel strip. 280 32 3300 800 Width (mm) Thickness (mm) Length Space 28 3.2 330 (mm) 80 (mm)

Engineering steel strip Type Thin sheet steel

Engineering steel strip Thin sheet steel

280 28

32 3.2

3300 330

800 80

Diameter (mm) 30 Diameter 3 (mm) 30 3

Approximately 28 days after casting, the roadside filling material used in the field test had a uniaxial 28 days casting, the filling material the field testsimilarity had a compressiveApproximately strength of 32 MPa (seeafter Figure 2) and anroadside elastic modulus of 11.2 used GPa.inBased on the uniaxial compressive strength of 32 MPa (see Figure 2) and an elastic modulus of 11.2 GPa. Based principle, the similarity ratio was found to have been 10:1. The components and proportions ofonthe the similarity principle, the similarity ratio was found to have been 10:1. The components and filling material are listed in Table 5. According to the constant of geometric similarity, standard cube proportions of the filling material are listed in Table 5. According to the constant of geometric specimens with widths, heights, and lengths of 200 mm were prepared using this material. Afterwards, similarity, standard cube specimens with widths, heights, and lengths of 200 mm were prepared the specimens were placedAfterwards, in a curing the boxspecimens at 25 ◦ C and a relative for°C28and days. using this material. were placed inhumidity a curing of box90% at 25 a relative To humidity investigate the shear performance of the backfill–truss support structure, the truss and filling of 90% for 28 days. body were combined in this To prepare backfill–truss support structure, holesand were To investigate theexperiment. shear performance of thethe backfill–truss support structure, the truss drilled in the filling body and steel strip. After the holes were cleaned, the anchoring agent was slowly filling body were combined in this experiment. To prepare the backfill–truss support structure, holes in the and bolts steel strip. After strip, the holes were cleaned, the anchoring agent was injectedwere into drilled the holes to filling fix thebody anchor and steel as illustrated in Figure 12 slowly injected into the holes to fix the anchor bolts and steel strip, as illustrated in Figure 12. Table 5. Proportions and parameters of different materials and physical characteristics of similar materials. Table 5. Proportions and parameters of different materials and physical characteristics of similar materials. Unit Weight Elastic Modulus I:B:S

G%

RA%

β

USC (MPa)

(kN/m3 )

(MPa)

I:B:S G% RA% Unit Weight (kN/m3) USC (MPa) Elastic Modulus (MPa) β 1:0.25:0.22 2.5 2.5 25% 25%5.0%5.0% 2.80 3.29 1119.66 1:0.25:0.22 2.80 3.29 1119.66 Remark: Remark: I, B, S represent iron powder, barite, and sand, respectively; G%, β, and RA% the total weight I, B, S represent iron powder, barite, and sand, respectively; G%, β, and RA% percentages the total of gypsum, rosin, and alcohol, respectively. weight percentages of gypsum, rosin, and alcohol, respectively. Steel strip

80mm

Filling body

150mm

Rock bolt B1

Steel strip Rock bolt B2 200mm

50mm

Rock bolt B2

80mm

100mm

200mm

Filling body

β

50mm

Rock bolt B1

20mm

200mm

(a)

(b)

(c) Figure 12. Direct shear test and size of backfill–truss support structure. (a) Front view; (b) side view;

Figure 12. Direct shear test and size of backfill–truss support structure. (a) Front view; (b) side view; (c) enlargement of local areas. (c) enlargement of local areas.

5.2. Experimental Schemes Pressure-shear testing was conducted on the filling body specimens (prepared according to the parameters listed in Table 1 and a curing age of 28 days) and backfill–truss support structure, using an RLJW-2000 rock-pressure testing machine (Chaoyang Test Instrument Co., Ltd., Changchun, China). By changing the inclination angles (β) of anchor bolts B1 and B2 in the truss support


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5.2. Experimental Schemes Pressure-shear testing was conducted on the filling body specimens (prepared according to the parameters listed in Table 1 and a curing age of 28 days) and backfill–truss support structure, using an RLJW-2000 Minerals 2017, 7, 75rock-pressure testing machine (Chaoyang Test Instrument Co., Ltd., Changchun, China). 14 of 18 By changing the inclination angles (β) of anchor bolts B1 and B2 in the truss support structure, five test structure, fivedesigned test schemes designed using β1), values of 90° 2), (scheme 1), 70°3),(scheme 2), 60° schemes were usingwere β values of 90◦ (scheme 70◦ (scheme 60◦ (scheme 45◦ (scheme 4), ◦ (scheme 3), 45° (scheme 4), and 30° (scheme 5). and 30 (scheme 5). InIneach eachscheme, scheme,(1) (1)the theshear shearforce forcewas was applied applied to to the the specimens specimens under under initial displacement displacement control thethe shear force reached 0.5 0.5 kN;kN; (2) a(2) normal stressstress of 5.12 control and andthe theloading loadingwas wasstopped stoppedwhen when shear force reached a normal of kN 40% of theofcompressive strength of theof filling body)body) was applied using the stress 5.12(approximately kN (approximately 40% the compressive strength the filling was applied using the control method; and (3)and Li (3) et al. the relationship between the advance velocity and stress control method; Li [27] et al.established [27] established the relationship between the advance velocity loading rate rate of the testing machine. According to to these findings, the and loading of the testing machine. According these findings, theshear shearforce forcewas wasapplied appliedto to the the specimens specimens at at aa cross-head cross-head speed speed of 0.25 mm/min mm/minuntil untilthe thespecimens specimenslost losttheir theirload-carrying load-carryingability. ability. By comparingthe theshear shear strengths of conventional the conventional specimens and backfill–truss support By comparing strengths of the specimens and backfill–truss support structure, structure, we analyzed the influences of the truss support structure and inclination angle of the we analyzed the influences of the truss support structure and inclination angle β of the anchor βbolts on anchor bolts on theofshear strength of the filling body. the shear strength the filling body. 5.3. Experimental Experimental Results Results 5.3. The experimental experimental results results are are shown shown in Figure 13. Compared Compared with with conventional conventional filling filling bodies, bodies, The the backfill–truss backfill–truss support support structure structure exhibited exhibited a larger ultimate shear strength. The The ultimate ultimate shear shear the strength of the backfill–truss support support structure structure increased with a decrease in the angle of the inclined inclined strength anchor bolts. bolts. When When the the anchor anchor bolt bolt was was tilted tilted at at an an angle angle of of 60°, 60◦ , its its shear shear strength strength reached reached 3.3 3.3 MPa, MPa, anchor which was 107.54% higher than that of the filling bodies without truss support structures. structures.

Shear strength of the backfill-truss support structure(MPa)

3.5

3.0

2.5

2.0

1.5 Conventional scheme 1 specimen

scheme 2

scheme 3

scheme 4

scheme 5

Figure Figure 13. 13. Results Results of of direct direct shear shear test test of of backfill–truss backfill–truss support support structure. structure.

6. Field Application 6. Field Application To verify the supporting effect of the backfill–truss support structure, field investigations were To verify the supporting effect of the backfill–truss support structure, field investigations were conducted on the No. 41101 main transportation roadway in the Awuzisu coal mine, Xinjiang (red conducted on the No. 41101 main transportation roadway in the Awuzisu coal mine, Xinjiang region, Figure 1). The specific geological and mining conditions of this roadway are shown above. (red region, Figure 1). The specific geological and mining conditions of this roadway are shown above. 6.1. Support Structure Structure 6.1. Parameters Parameters of of Backfill–Truss Backfill–Truss Support The The roof roof and and coal coal walls walls are are supported supported according accordingto tothe theparameters parametersset setfor forthe theoriginal originalroadway. roadway. Beside the roadway, cement-based paste was used to construct a truss support structure Beside the roadway, cement-based paste was used to construct a truss support structure with with the the filling wall being 1.5 m, 3.5 m, and 2.5 m thick in the upper, lower, and middle parts, respectively. Figure 14 shows the backfill–truss support structure. The proportions of this filling material are listed in Table 1. Left-handed high-strength anchor bolts with a yield load of 171 kN and tensile load of 266 kN, containing no longitudinal reinforcement, were used. The bolts were 20 mm in diameter and required a pre-tightening force greater than 60 kN. The inclined anchor bolts in the upper and


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filling wall being 1.5 m, 3.5 m, and 2.5 m thick in the upper, lower, and middle parts, respectively. Figure 14 shows the backfill–truss support structure. The proportions of this filling material are listed in Table 1. Left-handed high-strength anchor bolts with a yield load of 171 kN and tensile load of 266 kN, containing no longitudinal reinforcement, were used. The bolts were 20 mm in diameter and required a pre-tightening force greater than 60 kN. The inclined anchor bolts in the upper and lower parts of the filling body were 3 m in length. Both had an inclination angle of 60◦ , while the anchor bolt Minerals 2017, 7, 75 75 was 1.8 m long. Z2360 anchor agents with a length of 0.3 m were used15 15 ofanchor 18 2017, 7, 18 in theMinerals middle part toof the ends. In addition, a W-type steel strip with a width, length, and thickness of 220 mm, 2400 mm, to anchor anchor the the ends. ends. In In addition, addition, aa W-type W-type steel steel strip strip with with aa width, width, length, length, and and thickness thickness of of 220 220 mm, mm, to and 2.5 mm, used. was used. 2400 mm,respectively, and 2.5 mm, was respectively, 2400 mm, and 2.5 mm, respectively, was used.

(a) (a)

(b) (b)

Figure 14. 14. Backfill–truss Backfill–truss support support structure: structure: (a) (a) front front view view and and (b) (b) right right view. view. Figure

Figure 14. Backfill–truss support structure: (a) front view and (b) right view.

6.2. Construction Construction Technology Technology 6.2.

6.2. Construction Technology

The backfill–truss backfill–truss support support structure structure in in the the No. No. 41101 41101 main main transportation transportation roadway roadway was was The

established m behind behind thestructure workinginface face using the filling filling technique. The The roadway specific construction construction The backfill–truss support theusing No. 41101 main technique. transportation was established established 33 m the working the specific procedure was as follows: 3 m behind the working face using the filling technique. The specific construction procedure was as follows: procedure was as follows:

(1)

(1) Advance Advance support was undertaken undertaken on the working working face using two rows of of single single hydraulic hydraulic (1) was the using rows Advance supportsupport was undertaken on theon working face face using twotwo rows of single hydraulic props props erected erected in in the the intake intake airflow airflow roadway roadway and and return return airway airway of of the the working working face. face. These These props erected in the intake airflow roadway and return airway of the working face. These props were props were were arranged arranged at at row row and and column column separations separations2 of of 1000 1000 ×× 1000 1000 mm mm22,, and and the the arrangedprops at row and column separations of 1000 × 1000 mm , and the advance-support distance advance-support distance distance was was maintained maintained at at 25 25 m m (Figure (Figure 15). 15). After After the the supports supports of of the the advance-support was maintained 25 m (Figure 15). Afterremoved, the supports of theofroadway face-endprops were were normally roadwayat face-end were normally normally three rows rows single hydraulic hydraulic roadway face-end were removed, three of single props were removed, three rows hydraulic established insupport the roadway retained established in of thesingle roadway retainedprops in the thewere gob to to temporarily support the working working face.in the established in the roadway retained in gob temporarily the face. gob to temporarily theat working face. Thesein props fixed at columns, 1000 mmand intervals These props propssupport were fixed fixed at 1000 mm mm intervals in both were the rows rows and columns, and the in These were 1000 intervals both the and the supporting width was wasand 4000the mm. both the supporting rows and columns, supporting width was 4000 mm. width 4000 mm.

Advancing Support Support Advancing Shield support Shield support

Gob Gob Stake Stake Filling body body Filling gob-side entry entry gob-side retaining retaining

41101 longwall longwall face face 41101 Dam-boards Dam-boards A A

A A

60~80m 60~80m 280m 280m

25m 25m

(a) (a)

Figure 15. Cont.

Pipe Pipe

Pump Pump


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Dam-board

Immediate roof

Dam-board

Immediate roof

Plastic membrane

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Plastic membrane Dam-board

Filling body

Dam-board

Filling body Immediate floor

Immediate floor

Prop

(b)

(2)

(3)

(4) (5)

Prop Figure 15. Filling retained roadway along gob of(b) No. 41101 main transportation roadway: (a) plan Figure 15. Filling retained roadway along gob of No. 41101 main transportation roadway: (a) plan view and (b) cross-section A-A. viewFigure and (b) A-A. 15.cross-section Filling retained roadway along gob of No. 41101 main transportation roadway: (a) plan view and (b) cross-section A-A. (2) After the working face was advanced by 3 m, densely arranged wooden pillars were used

After theasworking faceforwas by 3in m, arranged pillars by were substitutes theadvanced single props thedensely side closest to thewooden gob, followed theused (2) After the working face was advanced by 3 m, densely arranged wooden pillars were used as substitutes for the single props in the side closest to the gob, followed by the construction construction of the filling frameworks. The lateral plate (2.8 m in height) was always as substitutes forthe theThe single props inwhile the m side closest towas the gob, followed by the the to of the filling frameworks. lateral plate (2.8 height) perpendicular perpendicular to horizontal plane, theininside plate wasalways perpendicular to construction of the filling frameworks. The lateral plate (2.8 m in height) was always inclined direction of the seam. Moreover, the distance between platesdirection in the floor the horizontal plane, while thecoal inside plate was perpendicular to the two inclined of the perpendicular to the horizontal plane, while the inside plate was perpendicular to the strata was always 3.5 m (see Figure 15). To maintain the stability of the filling frameworks coal seam. Moreover, the distance between two plates in the floor strata was always 3.5 m inclined direction of props the coal seam. the of distance between two plates the floor and roof, hydraulic were fixed on filling one side the lateral plate close to theinretained (see Figure 15). To maintain the stability ofMoreover, the frameworks and roof, hydraulic props were strata wasatalways 3.5 of m 1(see Figure 15). To maintain the stability of the filling frameworks roadway intervals m along the advancing direction of the working face. To prevent fixed on one side of the lateral plate close to the retained roadway at intervals of 1 m along the and roof,grout hydraulic props were fixed on aone sidediaphragm of the lateral the retained cement from escaping by seepage, plastic wasplate laid close on thetoinner side of advancing direction of the working face. To prevent cement grout from escaping by seepage, roadway intervals of 1 m along the advancing direction of the working face. To prevent the filing at plate. a plastic diaphragm was on thebyinner side ato ofplastic the plate. listed groutuniformly fromlaid escaping seepage, diaphragm was laid on the1,inner side of (3) cement After being mixed according the filing proportions in Table the filling After being uniformly according the proportions in Table 1, the materials materials were mixed transported to the to filling area throughlisted a pipeline beside the filling roadway to the filing plate. were touniformly the filling area through atopipeline beside the roadway construct the roadside filling body. (3)transported After being mixed according the proportions listed in Tableto1, construct the filling the (4) materials Approximately four to five days the area fillingthrough body was first constructed, roadsideto roadside filling body. were transported to theafter filling a pipeline beside thethe roadway filling body was supported according to the designed parameters of the truss. constructfour the roadside filling body. Approximately to five days after the filling body was first constructed, the roadside filling (5) Steps 1 to 4 were repeated entire was constructed. (4) Approximately four to five daysthe after thefilling fillingwall body was first constructed, the roadside body was supported according tountil the designed parameters of the truss. filling body was supported according to the designed parameters of the truss. Steps 1 to 4 were repeated until the entire filling wall was constructed. 6.3. Application Effect (5) Steps 1 to 4 were repeated until the entire filling wall was constructed.

After working 6.3. Application Effect face No. 41101 was advanced by 300 m, the remaining roadway along the gob

6.3. Application Effect began to be supported using the backfill–truss support. Observation stations III and IV were After working face No. 41101 advanced by behind 300 m, the roadway along the gob began respectively established at 41101 130was mwas andadvanced 180 m theremaining working face (see Figure 1a). The After working face No. by 300 m, the remaining roadway along the gob to bebegan supported the backfill–truss support. stations deformation (including convergence between theObservation roof and floor, as wellIII asand that IV between the to be using supported using the backfill–truss support. Observation stations IIIwere and respectively IV two were walls ofatthe roadway) of the rocks in the roadway was observed the stations using a established 130 m and 180 m surrounding behind the working face (see 1a). face Theat(see deformation (including respectively established at 130 m and 180 m behind theFigure working Figure 1a). The crossing between method, as shown in Figure monitoring results are presented Figure 16. the two convergence the roof and floor,6b. asThe well as theas two of the roadway) of the deformation (including convergence between thethat roofbetween and floor, wellwalls as in that between walls of the roadway) of the surrounding rocks at in the roadway was observed at the stations a in surrounding rocks in the roadway was observed stations using a crossing method, asusing shown 50 at the station III crossing method, as shown in Figure 6b. The monitoring results are presented in Figure 16. Figure 6b. The monitoring results areRoof-to-floor presented in Figure 16. Rib-to-rib at the station III Roof-to-floor at the station IV Rib-to-rib at the station IV

Convergence(mm) Convergence(mm)

40 50

Roof-to-floor at the station III Rib-to-rib at the station III Roof-to-floor at the station IV Rib-to-rib at the station IV

30 40

20 30

10 20

10 0

25

50 75 100 Distance behind 41101 longwall face (m)

Figure 16. Deformation of surrounding rocks in backfill–truss supported roadway. 0

25

50 75 100 Distance behind 41101 longwall face (m)

Figure 16. Deformation of surrounding rocks in backfill–truss supported roadway.

Figure 16. Deformation of surrounding rocks in backfill–truss supported roadway.


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When the backfill–truss support structure was applied, the deformation of the surrounding rocks gradually increased with the advance of the working face over distances of up to 40 m, with the maximum deformation of the surrounding rocks being 37 mm, as shown in Figure 16. The rate of deformation of the surrounding rocks gradually decreased when the working face was advanced by 50 m. Therefore, the truss support structure was found to improve the shear resistance of the cement-based filling body and the capacity to resist lateral deformation by integrating the filling body with the roof and floor strata. Consequently, we concluded that the roadside backfill–truss support structure could adapt to the large shear stresses caused by roof movement in the mining of steeply inclined coal seams. 7. Conclusions 1. After performing gob-side entry retention work for this working face, the roadside filling body in the gob-side entry retention area is expected to bear a large shear stress when the lateral main roof moves, sags, and compacts the rubble. In this stage, if the shear stress τ on the roadside filling body is larger than its ultimate shear strength τ U , shear failure is expected to occur in the filling body. 2. By establishing mechanical models of the gob-side entry retention in a steeply inclined coal seam, we obtained an equation for calculating the shear stress on the roadside filling body after the roof strata had stabilized. Based on field investigations and laboratory experiments, some parameters were obtained, including the force acting on the main roof strata from the coal wall and the relationship between the axial bearing capacity and compression ratio of the rubble. On that basis, the shear stress on the roadside filling body of working face No. 41101 was found to be 15.89 MPa. If only the cement-based filling body was used as the roadside support, the maximum shear strength of the filling body was 12.38 MPa, which failed to satisfy the design requirement for shear strength. 3. The roadside backfill–truss support structure proposed in this paper was comprised of three anchor bolts and a steel strip. Therein, two inclined anchor bolts were anchored to the roof and floor, while the third anchor bolt was fixed to the filling body to lie parallel to the coal seam. These three anchor bolts were connected using the steel strip. A laboratory shear test indicated that the ultimate shear strength of the structure was 107.54% higher than that of a cement-based filling body when the dip angle of the inclined bolt was 60◦ . In addition, field monitoring showed that roof-to-floor and rib-to-rib convergences only reached 42 mm when the working face advanced by up to 50 m when using the roadside backfill–truss support structure in situ. Acknowledgments: This study was supported by the National Natural Science Foundation of China (No. 51574154), the China Postdoctoral Science Foundation Funded Project (No. 2016M592220), and the Qingdao Postdoctoral Applied Research Project (No. 2015198). Postgraduate Innovative Fund Project of College of Mining and Safety Engineering (KYKC17005). The authors express their sincere thanks to the reviewers for their helpful comments and suggestions for improving this paper. Author Contributions: Jianguo Ning and Jun Wang developed the design solution and conducted the data analysis. Considering the contribution of Jianguo Ning and Jun Wang on the manuscript, they contributed equally to this paper and should be considered co-first authors. Tengteng Bu and Xuesheng Liu conducted the laboratory and analytical work in the interpretation of data. Shanchao Hu designed the laboratory work and revised the paper. Conflicts of Interest: The authors declare no conflict of interest.

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