study on the rock mass caving and surface subsidence...

Research ArticleStudy on the Rock Mass Caving and Surface SubsidenceMechanism Based on an In Situ Geological Investigation andNumerical Analysis

Fengyu Ren1 Dongjie Zhang 1 Jianli Cao1 Miao Yu 12 and Shaohua Li12

1Center of Mining Research School of Resources and Civil Engineering Northeastern University Shenyang 110819 China2Center for Rock Instability and Seismicity Research Northeastern University Shenyang 110819 China

Correspondence should be addressed to Dongjie Zhang 834381852qqcom

Received 16 April 2018 Revised 20 June 2018 Accepted 14 July 2018 Published 26 July 2018

Academic Editor Enrico Conte

Copyright copy 2018 Fengyu Ren et al This is an open access article distributed under the Creative Commons Attribution Licensewhich permits unrestricted use distribution and reproduction in any medium provided the original work is properly cited

To deeply understand the mechanism of the rock mass caving and associated surface subsidence during sublevel caving miningthe Xiaowanggou iron mine was selected as an engineering project case study The study area was analyzed by means of an in situgeological investigation and numerical simulation First a borehole television (BHTV) system and a GPS monitoring system wereused to monitor the caving process of the roof rock mass and the development of the surface subsidence the monitoring timewas thirteen months Then a numerical simulation was used to analyze the damage evolution of the rock mass Research showsthe following (1) in situ geological monitoring results indicate that the caving process of the roof rock mass presents intermittentcharacteristics where slow caving and sudden caving are conducted alternatively and the arched-caving trend is more pronouncedwith continuous mining The surface subsidence horizontal displacement and horizontal deformation of the hanging wall arehigher than that of the footwall and the subsidence center gradually deflects to the hanging wall in the late stage of the +45msublevel mining (2)Numerical simulation results indicate that the extension and penetration of the shear and tensile cracks alongthe joints and intact rock bridges are the main factors causing the rock mass caving and the existence of the stress arch and itsevolution process is the fundamental reason for the intermittent caving of the rock mass The rapid development of damage to thehanging wall (the damage angle reduced) is the main cause of the deflection of the subsidence center affected by joints and themining size (3) In the future of mining large-scale subsidence will occur on the surface of the hanging wall

1 Introduction

There are a large number of iron ore resources in the north-eastern part of China and the total reserves of these resourcesaccount for two-fifths of the national gross proved reservesDue to the influence of geological conditions orebody distri-bution characteristics and mining technology most of theseunderground iron mines are mined by the sublevel cavingmethod [1ndash3] which is characterized by extraction on a largescale and high production efficiency with low mining costHowever sublevel caving mining inevitably results in theformation of underground goafs [4] and the existence ofunderground goafs will cause the imbalance and redistri-bution of the stress around the opening which will causethe displacement deformation and failure of the roof strata

[5 6] More seriously the failure will cause the roof stratacaving to the goaf greatly threatening the mine operationsand personnel safety [7 8] The caving process propagatesupwards to the surface thus significant surface subsidenceappears and even large-scale collapse pits are formed [9] As aresult this poses a serious threat to buildings haulage roadsrivers industrial facilities and farmland [10ndash12] To controlthe dangerous situation it is of great importance to study themechanism of the rock mass caving and associated surfacesubsidence

Rock mass caving is the product of a complex rock massresponse to excavation [13] and rock failure is the mostimportant factor causing the occurrence of a caving Becauseof the complexity of the underground rock distribution thepartial failure of the rock will lead to a series of chain failure

HindawiMathematical Problems in EngineeringVolume 2018 Article ID 6054145 18 pageshttpsdoiorg10115520186054145

2 Mathematical Problems in Engineering

[6] and thus induce massive caving of the roof strata Theformation of surface subsidence is the inevitable result of thecaving to the surface

The rock mass failure and surface subsidence caused byundergroundmining depend onmany factors such asminingdepth geological conditions and rock mass structure [14ndash16] Many studies have conducted in-depth analysis of themining-induced rock mass failure and surface subsidencewhich can be divided into three categories The first categoryis the physical similarity simulation method By establishinga model similar to the actual engineering the displacementstress and failure under the influence of mining can beanalyzed Ju et al [17] studied the surface stepped subsidencemechanism and proposed the important role of primarykey stratum and subkey stratum in the development ofsubsidence Sun et al [18] studied the characteristics ofroof strata movement and the evolution of water-conductingfractures and proposed SBM technology to effectively con-trol strata movement and deformation Huang et al [19]studied the characteristics of the goaf-roof displacement andstrain through physical similarity experiments and reason-ably determined the key parameters required for miningdesign The second category is on-site monitoring methodsuch as total station surveys GPS field surveys differentialInSAR techniques borehole monitoring and microseismicmonitoring Xia et al [20] used borehole monitoring tech-nology to analyze the caving process of the goaf-roof andpointed out that the caving evolution was affected by thestructural character of the rock mass Wang et al [21] used acombination of on-sitemonitoring and numerical simulationto analyze the fracture development of the rock mass andsurface subsidence characteristics near the riverbed andproposed the Four-Element Structure of antiseepage for theriverbed Zhao et al [22] used microseismic technology toanalyze the stability of goafs and slope during the transitionfrom open-pit to underground mining and to study the rockfailure mechanism The third category is numerical simula-tionmethods such as the finite elementmethod (FEM) finitedifference method (FDM) and discrete element method(DEM) which have developed rapidly in recent years Zhao etal [23] analyzed the rock movement mechanism and surfacesubsidence characteristics under different stress fields Xuet al [24] proposed an EJRM-based numerical method tostudy the strata and surface movement under the influenceof a jointed rock mass He et al [25] used a discontinuousdeformation analysis to study the formation and stabilizationof a stress arch in laminated rock mass

Based on the above analysis results each method has itsown advantages However the physical similarity simulationmethod is mainly focused on underground coal mining andthe rock structure of metal mines is different from that ofcoal mines [23] The on-site monitoring method is mainlyused to evaluate and analyze the trend of rock caving andsurface subsidence but cracks and stress evolution in therock mass cannot be represented visually When we usenumerical methods for analysis the simulation results maynot be consistent with actual site conditions [26 27] Theneeds of accurate study of the mechanism of caving andsubsidence cannot bemet solely using onemethod and there

are relatively few studies on the cavingmechanism of the rockmass in metal mines Therefore using a combination of on-site monitoring and numerical simulation it is possible torealize a comprehensive explanation of the sublevel cavingmining from the macroscopic rock mass collapse modeland surface subsidence trend of the caving mechanism ofroof strata and the microscopic crack propagation law Thisconnects mining and rock failure in time and space thusproviding support for mining safety and the protection ofsurface environment and facilities

In this paper the Xiaowanggou iron mine was selected asan engineering project for an in-depth study The rock masscaving mechanism and surface subsidence characteristicsduring the sublevel caving mining were comprehensivelystudied using an in situ geological investigation and numer-ical analysis method (the FEM-based numerical softwareRFPA 2D [28ndash31]) The overall understanding of the processfrom the crack propagation stress evolution and damagedevelopment to the rock caving and surface subsidence isobtained The research can provide guidance for other metalmines with similar conditions

2 Background of Geologic andMining Conditions

The Xiaowanggou iron mine is a sedimentary metamorphicmagnetite deposit and is located in Liaoyang City LiaoningProvince northeast ChinaThe mining area is approximately70 km long and the elevation is approximately 250m It hasa temperate continental monsoon climate with an averageannual precipitation of approximately 800mmThe ore bodyis stratoid and the strike is north by east 50∘ with the dipdirection of south-east and the dip angle ranging from 30∘ to50∘ The strike length is approximately 350m and the averagethickness is 100m Thus it is a typically inclined metal minethe deep mining area extends from 190 m below the surfaceto over 340 m and the surrounding rock is mainly mixedgranite

The Xiaowanggou iron mine is mined by partition Theentire deposit from top to bottom is divided into threeminingareas namely the uppermining area themiddlemining areaand the deep mining area and there is a safety pillar betweenthe partitions [32] The deep mining area was selected as theresearch object of this studyThe sublevel caving method wasused in this mining area and the first mining section waslocated at the +60m level with a subsection height of 15mAt present the major mining activities are carried out at the+30m level and the exploration system has been arrangedto the 0 m level In the mining process the overlying strataface continuous fracturing and caving In the late stage of+45m sublevel mining an oval collapse pit with a maximumdepth of 27m and a diameter of about 80mwas formed on theground in October 2015 Multiple fracture lines were formedaround the collapse pit which poses a serious threat to theenvironment surface industrial facilities and haulage roads(Figure 1)

This is of great significance for grasping the process ofrock mass caving and the trend of surface subsidence inadvance which can effectively reduce the potential safety

Mathematical Problems in Engineering 3

ChinaBeijing

Liaoyang

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

4

2480

2423

2550

2574

1

2 3

4

50m

C1

C3

C2

C6

C4

C5

Collapse pit boundary

Outline of surface cracks

Ore body boundary

Observation cracks

Monitoring station

Monitoring borehole

Haulage road

Exploration line

05m

Collapse pit

27m

Cracks

80m

Road

Collapse pit

Crack

Cracks

North

Figure 1 The main research area in the Xiaowanggou iron mine

hazard caused by mining on surface roads and facili-ties

3 Analysis of the In Situ GeologicalInvestigation

31 In Situ Observations of Rock Mass Caving in BoreholesThe mining of the +45 m section started in May 2014

In order to monitor the roof rock caving situation inreal-time the mine had four geological monitoring boreholes arranged along the 3 exploration line (Figure 2) tobe used in combination with the geological explorationresults The average spacing of the boreholes was 30 mcovering the main subsidence area affected by the miningactivities The relevant technical parameters are shown inTable 1


Table 1 Borehole data for roof monitoring

Borehole number Elevation Borehole depth(m)Surface (m) asl Bottom (m) asl

1 2391 891 1502 2325 825 1503 2341 841 1504 2363 863 150

1st mining layer

Monitoring borehole

Ore body boundary

2nd mining layer

3rd mining layer

Mixed granite

2150 2250 2350 24500m

60m

100m

200m

250mE W

247m

0m

100m

45m

15m30m

1 2 3 4

Unexploited ore body

Figure 2 The geological profile of exploratory line No 3

Monitoring was performed beginning with the comple-tion of the geological borehole project The BHTV system ismainly used to monitor the caving process of the roof rockmass and it is also widely used to obtain statistics about jointsand fissures [4] The monitoring time was thirteen monthsand the initial monitoring was carried out once per monthIn the late stage of monitoring with the acceleration of thecaving of the rock mass the monitoring time interval wasadjusted When the roof rock mass in the 2 borehole wasonly approximately 30 m away from the ground surface thepersonnel and equipment were evacuated immediately OnOctober 5 2015 a massive collapse occurred on the groundsurface

Scanning images in Figure 3 show that a clear fracturezone appears at a certain distance above the bottom of theborehole It is noteworthy that the fracture position of the twoboreholes differs by 73 m this may be related to the size effectof the rock mass structure [33] Through regular monitoringaccurate data of the roof caving process are obtained and themonitoring analysis results are shown in Figure 4

In the early stage of monitoring the stability of theroof mixed granite was good As the mining advanced the

goaf-roof was always in the slow caving stage (five months)As themining span continued to increase the first large-scalecollapse took place between January and February 2015 andthe caving curve was an obvious arch (Figure 4(b)) Duringthe entire monitoring process a total of three sudden cavingperiod took place when the caving heights were 152 m236 m and 481 m (Figure 4(a)) The entire caving projectcould be described as alternating development with slowsudden caving In the later monitoring period the roof stratacollapsed rapidly and caved to the surface in October 2015The total height of the caving was approximately 80 m andthe time to collapse to the surface was only 20 days indicatingthat the roof strata activity in this stage was more intense

32 Monitoring of SurfaceMovement and Subsidence Surfacesubsidence is the inevitable result of the rock caving to thesurface [34] In order to ascertain the range and degree ofsurface rock movement and subsidence caused by miningGPS monitoring is conducted in the mining influenced area

According to the geological and mining conditions asurface monitoring system consisting of two monitoringlines was established along the strike and inclination of orebodies As shown in Figure 5 fifteen monitoring points (R1R2 R3 R15) were arranged along the strike and fourteenmonitoring points (N1 N2 N3 N14) were arranged alongthe inclination and the interval between monitoring pointswas 20m These monitoring points were controlled by sixmonitoring stations (C1 C2 C3 C6)

The recovery of the +45m sublevel ore body began inMay 2014 and was finished by the end of July 2015 with atotal recovery time of approximately fourteen months andsurface subsidence monitoring was carried out monthly Thehaulage road is located within the range of rock movementaffected by mining therefore it is crucial to carry out riskassessment of the haulage road Since the influence of surfacerock movement along the inclination is significant for thehaulage road the study will focus on the analysis of thecharacteristics of the surface deformation and subsidencealong the inclination

Themain characteristics of surface subsidence and defor-mation are summarized as follows(1) With continuous mining of the +45m sublevel

ore body the mining-induced surface subsidence graduallyincreased The lowest point of subsidence mainly appearedin the center of the ore body and the subsidence rateincreased as shown in Figure 6(a) After January 2015 thesubsidence center slowly migrated to the hanging wall andthe maximum subsidence rate decreased By comparing the


53 m

Collapse boundary ofthe bottom borehole

Fracture

(a)

126 m


Fracture

(b)

Figure 3 Borehole scanning images based on the BHTV system on April 17 2015 (a) 2 borehole and (b) 3 borehole

159

25

159

20

159

24

159

14

156

18

155

15

152

17

148

15

151

16

Time (YearMonthDay)1 borehole2 borehole

3 borehole4 borehole

1105

624

589

353245

93

481m

236m

152Suddenly caving

Slowly caving

0

20

40

60

80

100

120

Roof

rock

cavi

ng h

eigh

t (m

)

(a)

431 2Borehole number

140815140916141013141115141217150116150217

150315150417150515150618150716150815150914150920150924150925

Caving arch

Middle caving lineEarly caving line

Initial caving line

Late caving line

0

20

40

60

80

100

120

140

Roof

rock

cavi

ng h

eigh

t (m

)

(b)

Figure 4 BHTV monitoring results (a) changes in the caving height of the roof strata at different boreholes from Aug 15 2014 to Sept 252015 (b) caving monitoring curves at the four boreholes from Aug 15 2014 to Sept 25 2015


2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325

Ore body boundaryof +45m level

Different periods ofmining progress

Mining direction

North

4Haulage road

Figure 5 Plane view of the distribution of monitoring points

far-field subsidence value the largest cumulative subsidencein the hanging wall was clearly higher than that in thefootwall By the end of July 2015 the maximum subsidencecaused by the +45m sublevel mining was of 10133 mmwhile the maximum subsidence near the haulage road wasapproximately 325 mm(2)The maximum horizontal displacement appeared in

the vicinity of the hanging wall and footwall boundaries ofthe ore body not the center On the contrary the horizontaldisplacement at the center of the ore body was almostzero with the largest subsidence The maximum horizontaldisplacement in the hanging wall was about 2 times that inthe footwall as shown in Figure 6(b) During the miningprocess the zero point of horizontal displacement movedslowly towards the hanging wall By the end of extractionthe zero point of horizontal displacement moved to the rightby approximately 114m and the maximum horizontal dis-placement was -2857mm in the hanging wallThemaximumhorizontal displacement near the haulage road was about278m where the negative value indicated the horizontaldisplacement in the hanging wall and the positive valueindicated the horizontal displacement in the footwall(3) The surface horizontal deformation mainly showed

wave-shaped changes along the inclination As the miningadvanced the surface horizontal deformation in the far-fieldmainly alternated with tensile and compressive deformationNear themining area the horizontal deformation wasmainlymanifested as tensile deformation and increased graduallyas shown in Figure 6(c) The horizontal deformation near

the center of the ore body decreased and it was worthnoting that the horizontal deformation in the hanging wallwas significantly higher than that in the footwall and themaximum horizontal deformation in the hanging wall wasapproximately 14sim19 times that in the footwall Near thehaulage road the maximum tensile deformation was 095mmmand themaximumcompressive deformationwas -053mmm and several cracks had appeared on the pavementrecently (Figure 7) Positive values represented horizontaltensile deformation and negative values represented horizon-tal compression deformation(4) Based on the monitoring results of monitoring point

N8 the dynamic subsidence of the surfacewasmainly dividedinto three stages the slow subsidence period the activesubsidence period and the attenuation subsidence period asshown in Figure 6(d)When themining operation reached 28m it entered the active subsidence period and the subsidenceand subsidence rate increased significantly The severe subsi-dence occurred in the range of 42sim64 m and the maximumsubsidence rate reached 107 mmd The active subsidenceperiod under the mining-influence had a long duration ofapproximately 193 days and the subsidence reached 915When the mining had advanced to 80 m the subsidenceand subsidence rate tended to be stable and the attenuationsubsidence period was entered It was noteworthy that themaximum subsidence rate did not represent the largestsubsidence

Based on the distribution characteristics of the curves ofsubsidence (similar to s-type) and subsidence rate (normal


Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720

+45m mining areaHaulage road

minus1200minus1100minus1000

minus900minus800minus700minus600minus500minus400minus300minus200minus100

0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road

minus400minus350minus300minus250minus200minus150minus100

minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720

minus4minus3minus2minus1

0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)

20 40 60 80 1000Mining distance (m)

0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)

Monitoring subsidence rateMonitoring subsidence value

(d)

Figure 6Monitoring results of the ground subsidence and deformation (a) surface subsidence at differentmonitoring periods (b) horizontaldisplacement at different monitoring periods (c) Horizontal deformation at different monitoring periods (d) subsidence rate and value ofmonitoring point N8 at different mining distance

distribution) in Figure 6(d) a dynamic subsidence functionexpression is established

119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)

where D0 is the final settlement value of the surface at themonitoring point mm x is the horizontal distance betweenthemonitoring point and themining facemDx is the surfacesettlement value at the monitoring point when the distance isx mm l is the distance between themonitoring point and themining open-cutm v is themining velocitymdmandn arethe fitting coefficients When mining operations are carried

out at a constant velocity v the surface settlement rate at themonitoring point is expressed as follows

V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)

sdot [1 minus exp(minus119898 (119909 + 119897)V)]119899minus1

(2)

According to (1) and (2) the curves of subsidence andsubsidence rate are shown in the solid line in Figure 6(d)


Surface cracks

Figure 7 Cracks on the haulage road

where the fitting curve basically agrees with the monitoringresult The above expression can provide guidance for thedynamic prediction of the process of surface subsidence

4 Modeling Work

Through on-site monitoring the macroscopic process ofrock mass caving and the development trend of surfacesubsidence can be clearly observed However it is difficultto accurately explain the mechanism of rock mass cavingand surface subsidence through on-site monitoring In orderto further study the mechanism of rock mass caving andsurface subsidence in the Xiaowanggou iron mine RFPA 2Dnumerical software is used in this study

41 The Principle of RFPA 2D RFPA 2D can be used to sim-ulate the nonlinear deformation and discontinuous mediummechanics of quasi-brittle rocks The material properties ofheterogeneous rocks [35 36] including Youngs modulusPoissons ratio and intensity properties are randomly dis-tributed in the entire analysis domain Assume these elementsare subject to the Weibulls distribution [37] defined asfollows

120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)

where 120583 is the parameter of the element (such as theYoungrsquos modulus Poissonrsquos ratio or strength properties) 1205830is the average of the element parameter m is a parameterdefined by the shape of the distribution function and repre-sents the degree of material heterogeneity

In RFPA 2D the maximum tensile strain and Mohr-Coulomb criterion are used to define the damage thresholdthe former is used to determine whether the element issubject to tensile damage and the latter is used to determinewhether the element is subject to shear damage The elasticmodulus of the damaged material is defined as follows

119864 = (1 minus 119863) sdot 1198640 (4)

where D is the damage variable and E and E0 are theelastic modulus of the damaged and undamaged elementsrespectively

tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)

Figure 8 Elastic-brittle damage constitutive law of elements sub-jected to uniaxial stress (a) The case under uniaxial tensile stress(b) the case under uniaxial compressive stress [31]

Figure 8 presents the elastic-brittle damage constitutiverelation with a given specific residual strength When thetensile stress in the element reaches the tensile strength 1198911199050ie 1205903 le minus1198911199050 the damage variable D can be defined as

119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)

where 120582 is the residual tensile strength coefficient 119891119905119903 =120582 sdot 1198911199050 and 119891119905119903 is residual tensile strength 1205761199050 is the strainat the elastic limit and 120576119905119906 is the ultimate tensile strain atwhich the element completely damaged in tension as shownin Figure 8(a) Ultimate tensile strain is given as 120576119905119906 = 120578 sdot 1205761199050120578 is the ultimate strain coefficient

TheMohr-Coulomb criterion is used as a second damagecriterion to describe the element damage under compressivestress conditions (Figure 8(b)) and the expression is asfollows

1205901 minus 1 + sin1205931 minus sin1205931205903 ge 1198911198880 (6)

where 1205901 and 1205903 are the maximum and minimum prin-cipal stress respectively 120593 is the friction angle 1198911198880 is the


Table 2 Parameters of the major joint sets

Joint set Orientation(∘) Spacing(m)Dip direction Dip Maximum Minimum Average

1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122

Sub-horizontal joint

Steeply dipping jointsSteeply dipping joints

Figure 9 Geological structure of the rock mass

uniaxial compressive strength The damage variable undercompression is described as follows

119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)

where 1205761198880 is the compressive strain of the elastic limit 120582 isthe residual strength coefficient and 120582 = 1198911198881199031198911198880 = 1198911199051199031198911199050 isassumed to be true when the element is under compressionor tension

In numerical analysis when the element is damagedits stiffness and strength are reduced and the model isthen recalculated under the new parameters The next loadincrement is added only when there are no more elementsstrained beyond the strength-threshold corresponding to theequilibrium stress field and a compatible strain field

When using RFPA2D for numerical simulation blackcracks will appear in the postprocessing graph after theelement exceeds the tensile limit The simulation of thecrack is similar to the fuzzy crack model ie the crack issmeared over the whole element which greatly simplifies thesimulations of crack generation expansion and propagation

42 Determination of Mechanical Parameters in NumericalModel According to the geological survey results the roofstrata mainly consist of mixed granite and there is nomajor fault-broken zone in the study area The rock mass isobviously influenced by gneiss structure it is the main factorin the formation of joints (Figure 9) The joints are mainlycomposed of tectonic and weathered fractures The existenceof joints has a significant influence on the development ofthe caving and the characteristics of surface rock movement

[24 38ndash40] which usually leads to a large number of unstablerock blocks in the formation The detailed survey of thestructural plane of the roof rockmasswas carried out byman-ual in underground roadways the main survey parametersincluded occurrence inclination dips and spacing of jointsand the measured parameters were then analyzed usingDIPS software [4] Ultimately we identified the dominantstructural planes Based on geological survey results thereare mainly three joint sets in the rock mass Figure 10 showsthe distribution of joint sets in the study area The relevantparameters of the joint sets are shown in Table 2

As shown in Figure 10 and Table 2 the optimal jointsets obtained from the calculation and analysis consist ofa gently dipping joint and two almost orthogonal steeplydipping joints this rock mass structure is the most favor-able for collapse [41] For simplicity and convenience two-dimensional numerical models are constructed by appropri-ately simplifying the three joint sets with ang0∘ ang70∘ and ang80∘The mechanical parameters of the roof mixed granite aremainly obtained through laboratory tests [42 43] as shownin Table 3 The elastic modulus of the rock mass is calculatedusing formula (8) [24]

119864 = 10(119877119872119877minus10)40 (8)

where E is the elastic modulus of the rock mass and RMRis the geomechanics classification index of the rock mass [4445]

When the in situ geological model is used for numericalsimulation it is very important to correctly choose the inputparameters but also the correct boundary conditions anddimensions of the model which directly determine thereliability of the simulation results In combination with theactual surface subsidence range obtained in the field the


Figure 10 The distribution of the major joint sets in the study area

Table 3 Mechanical parameters of the mixed granite

Rock type

Mechanical parameters

ElasticmodulusE [GPa]

Uniaxialcompressivestrengthf c0 [MPa]

PoissonrsquosratioV

FrictionangleΦ [∘]

Density119875 [kg mminus3]

Mixedgranite 35 553 024 30 2700

model dimensions should be greater than this range andthe boundary conditions can refer to relevant references[28 29 31] Here we highlight how to determine thephysical-mechanical parameters used in numerical simula-tion through the measured physical-mechanical parametersof the rock mass and the homogeneity coefficient

In order to solve the practical problems a numericalmodel is first established Its geometrical dimensions areon the same order of magnitude as the actual engineeringproblems to be studied The numerical model size is 200mtimes100 m and is divided into 20000 finite element grids Inthis paper the coefficient of homogeneity of the sample isselected as m=3 and the relevant parameters input in themodel are selected according to the fitting formulas (9) and(10) [46]

1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)

where E0 and f c0 represent the elastic modulus andintensity values calculated in the model respectively E andf c represent the macroscopic elastic modulus and intensityvalues of the model (measured values) respectively

As shown in Table 3 the compressive strength (f c) of themixed granite is 553 MPa and the elastic modulus (E) is35 GPa The calculation inputs for the compressive strength

and elastic modulus are calculated according to formulas (9)and (10) the values are 689 MPa and 113 GPa respectivelySubsequently the uniaxial compression numerical simula-tion of the model is carried out and we observe the peakintensity of the model Because the numerical calculation hasgood repeatability and maneuverability the initial value ofthe compressive strength and elastic modulus can be adjustedwithin a narrow range until the compressive strength of themodel obtained is consistent with the rock mass strengthof the actual engineering As shown in Figure 11 the peakstrength of the model has reached 549 MPa which is veryclose to the measured compressive strength of the mixedgranite

The average length and spacing of the joints based on thegeological survey results are approximately 25 m and 15 mrespectively It should be noted that the joints used in themodel are assumed to be made of a ldquoweak materialrdquo withlower strength and stiffness Wong et al [47] reported thatthe homogeneity index should be greater than 20 but then fallwithin the typical range of 20ndash60 Therefore m=3 is chosento characterize the heterogeneity of the rockmass Finally wedetermine the input parameters of the numerical model asshown in Table 4

43 Numerical Modeling In this study the geological profileof exploratory line No 3 is selected and modeled for thesubsequent numerical simulation (Figure 2 minor surfaceundulations and contact zones are ignored) The numerical


Table 4 Parameters used in the numerical model

Rock type


Homogeneityindex m



Uniaxialtensilestrengthf t0 [MPa]

PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)

Figure 11 The mixed granite strength curve obtained by numericalcalculation

model is shown in Figure 12 the analysis domain is 600times250m and the model is divided into a mesh with 150000elements

The top boundary is set as free and normal displacementsare constrained on the right left and bottom boundaries themodel is affected by its own gravity The draw is simulated bygradually removing the rock mass at a depth of 175 m (the1st mining layer) and for each step the excavation size is15 m in height and 10 m in length The model is calculatedin a quasistatic manner to achieve a balanced state and thecalculation process is assumed to be a plane strain problemThen we will describe the progression of rock mass cavingand associated surface subsidence through the excavationwith an advance increment of 10 mstep

In the caving process fracture initiation always starts fromthe joint sets and propagates into the rock bridges [48] Inorder to highlight the role of rock bridges in the caving ofrock masses we assigned a persistence lt1 to the joint setsand the nonpenetrative area between joints in the numericalmodel is considered as the rock bridges The uneven spatialdistribution of the structural planes in the rock mass mustinevitably lead to uneven distribution of length and positionof rock bridges [49 50] Affected by the existence of jointsrock bridges are generally randomly distributed in the rockstrata Therefore in numerical models randomly distributedrock bridges were used for simulation and the rock bridgeswere not quantified in detail

5 Numerical Simulation Results

51 RockMass Caving Figure 13 shows the numerical resultswhere the crack failure process stress distribution anddamage development can be more intuitively demonstratedthroughout the numerical simulation process In the initialstage of excavation the initial balance within the rockmass isdisturbed causing the stress to be concentrated on both sidesof the undercut (the shading intensity indicates the relativemagnitude of the shear stress) At the same time the goaf-roof produces tensile and shear cracks along the joints underself-weight In the damage development graph the red cellsrepresent tensile damage and the white cells represent sheardamage (Figure 13(a)) As the excavation continues whenthe advancing distance reaches the ultimate caving span [51]tensile and shear cracks interconnect along the joints andintact rock bridges and the roof strata begin to collapse inthe goaf until it forms a self-stabilizing structure Tensiledamage occurs at the bottom of the roof strata and the sheardamage continues to expand upward along the joints in thisprocess However the rock mass on both side walls showsdifferent caving mechanisms The sidewall rock mass at thebottom is mainly sliding failure which is conducive to thedevelopment of failures towards the footwall It graduallyturns into a toppling failure with the continuous cavingwhich is conducive to the development of failures towardsthe hanging wall (Figure 13(c)) It is worth noting that oneor more obvious stress arches [52 53] are formed above theself-stabilizing roof strata (the zone presented with brightercolor) and the stress is mainly concentrated within the stressarch while the stresses below are diminished The rock masscaving is mainly due to the tensile failure The presence ofstress arches limits the further collapse of the roof strataand it also causes the rock mass to exhibit an arched-cavingcharacteristicsThe roof strata entered a slowly caving periodat this time as shown in Figure 13(b) The stress arch mustbe broken for mining to continue this will force the supportpoints of the stress arch to move upwards and with thecontinuous caving of the rock mass below the stress archit enters the suddenly caving period When the caving lineapproaches the new stress arch it once again enters the slowlycaving period and the arched-caving trend is more obviousMeanwhile there has been a small subsidence on the surfacebut the roof strata have not completely collapsed and therocks of the side wall remain in a stable state (Figure 13(d))In the whole mining process the collapse rate of roof rockmainly presents periodic characteristics It can be expressedas alternating development with slowly suddenly collapse


250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body

Footwall Hanging wall

600m

Figure 12 Model setup for numerical modeling

and eventually surface subsidence This caving mechanism isbasically consistent with the on-site monitoring results of theXiaowanggou iron mine

52 Surface Subsidence The numerical results of the +45 msublevel excavation are shown in Figure 14 When the roofrock caves to the surface the stress arch has completelydisappeared and forms a significant subsidence trough onthe surface the tensile and shear damage has fully developedto the surface and the subsidence is large in the centralarea and decreases progressively towards the sidewalls At thesame time we observe that the stress concentration in thehanging wall is more pronounced than that in the footwalland in turn accelerates the development of destruction andcaving of the hanging wall rock mass surface disturbanceswithin this zone are caused by changes in the stress field andthe damage development has fully extended to the surfaceand continues to extend to the sidewall (Figure 14(a)) Withfurther excavation the rock mass in the nonsteady area ofthe hanging wall is separated and destroyed which resultedin a corresponding increase of the surface subsidence rangeand internal damage development (Figure 14(b)) When theexcavation distance reached 80 m relatively large topplingfailure of the rock mass on both sidewalls has occurred alongthe joints (Figure 14(c)) and these joints are more conduciveto the damage development of the hanging wall rock massAfter +45 m sublevel excavation was completed the rockmass on both sidewalls has been destroyed and subsidedcompletely and the damage angle of each sidewall has furtherdecreased (Figure 14(d)) In this study the damage angle isobtained based on the development of tensile damage

The numerical results of surface subsidence for the +45m sublevel at different excavation distances are shown inFigure 15 The depth of the subsidence trough increases asthe excavation distance increases The maximum subsidencevalue is produced in the mining center area subsidencedecreases as the distance from the central area increasesEventually the center of the subsidence trough deflectstowards the hanging wall and the asymmetrical featureis remarkable The concept of the subsidence trough was

gradually used to assess the degree of surface subsidencecaused by mining this concept takes into account one ofthe most important observed facts about subsidence namelythat the surface area affected is larger than the mined area[54] There is a certain deviation between the maximumsubsidence values obtained by numerical and on-site mon-itoring with the main reason being that the bulking effectof the collapsed rock mass is not taken into account in thenumerical analysis which may lead to a large value in thenumerical monitoring However the numerically obtainedcharacteristics of the surface subsidence are basically thesame as the on-site monitoring as already mentioned whichcan better explain the on-site surface subsidence mechanism

6 Analysis and Discussion

In this paper the Xiaowanggou iron mine is taken as anengineering case study and the combination of on-sitemonitoring and numerical simulation is used to study themechanism of the rock mass caving and related surfacesubsidence in sublevel caving The process of rock masscaving and the development trend of surface subsidence areobtained through an in situ geological survey The processesof crack stress and damage evolution in the rock massduring mining are obtained through a numerical inversionThis can be used to recognize the whole failure processfrom the beginning of the caving to the formation of surfacesubsidence allowing for an intuitive understanding of therock mass caving and related surface subsidence mechanismThe research results can provide guidance for other mineswith similar mining conditions

According to the in situ geological survey the arched-caving development of the rock mass has been verified Theentire caving process can be described as alternating devel-opment with slow sudden caving it is concluded that thistype of development is related to the arch effect of rock masscaving The effect of mining the +45 m sublevel ore body onsurface subsidence and horizontal deformation is significantAlthough the subsidence rate of the hanging wall is largerthan that of the footwall the horizontal deformation near the


Cracks initiation

Stress concentration area

72∘ 76∘

(a)

Stress arch Damage angle(Footwall)

Damage angle(Hanging wall)Caving arch

70∘ 73∘

(b)

Stress arch Toppling failure

Sliding failure

68∘ 71∘

(c)

Failure processStress distribution Damage development

Stress arch

Caving arch

66∘ 67∘

(d)

Figure 13 Numerical simulation results of the +60 m sublevel with different advancing distances (a) 40 m (b) 60 m (c) 80 m (d) 100 m


Damage developmentUnstable areaStress concentration

64∘62∘

(a)

Stress concentration

59∘61∘

(b)

Toppling failureStress concentration

59∘54∘

(c)

Stress distribution Failure process Damage development

56∘52∘

(d)

Figure 14 Numerical simulation results of +45m section with different advancing distances (a) 40m (b) 60m (c) 80m (d) 100m


Distance (m)

advancing distance=20madvancing distance=40m


advancing distance=100m

-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area

Figure 15 Numerical results of surface subsidence with differentadvancing distance (+45m sublevel)

haulage road reached 095 mmm by the end ofMay 2015 andsome cracks had been observed in the fieldThere is no doubtthat the mining of the +30 m and +15 m sublevel ore bodieswill increase the risk of road breakage and collapse

In the analysis of rock mass caving the crack failureoccurs mainly along the randomly distributed joints (weaklyconstructed) and extends into the intact rock bridges as thefailure path of rock mass is not straight but flexural Theroof strata will not collapse to the goaf until the shear andtensile cracks have completely penetrated Once the intensityof the stress concentration in a caved zone of the roof stratais released a stress shadow area (the zone presented withgreyer color) and an apparent stress arch (the zone presentedwith brighter color) above the caved zone will be producedas shown in Figure 13 (the stress distribution) When oneor more key rock bridges are damaged the stress arch willbreak immediately With the increase in advancing distancethe stress arch is repeatedly formed and dissipated andcontinues to develop upward Finally the roof strata collapseto the surface to form a significant subsidence trough withthe destruction of the last stress arch Understanding thecaving mechanism of the rock mass during the mining is anessential aspect for accurate prediction of the rock fracturepropagation and the surface subsidence development it canaid in the timely adjustment of the mining sequence putforward preventive measures and provide an effective guar-antee for mine safety exploitation and surface environmentand structural protection

In the analysis of ground subsidence the concept of thedamage angle is introduced The curve of the damage anglewith different excavation distances is shown in Figure 16all data are obtained from the numerical models and eachpoint corresponds to a different model where damage angles

+45m sublevel+60m sublevel

Inflection point

The subsidencecenter is biased toward

the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)

20 40 60 80 100 120 140 160 180 200 2200Advancing distance (m)

Hanging wallFootwall

Figure 16 Variations of damage angles with different advancingdistance

decrease with an increasing excavation distance It is worthnoting that in the 1st mining layer (+60 m sublevel) thedamage angle of the hanging wall is always larger than thatof the footwall and the angle difference decreases as theadvancing distance increases indicating that the rock massin the footwall is affected greatly in this stage With theincrease in excavation distance this effect is graduallyreduced Because of the existence of the stress arch thereis no significant subsidence on the surface After the orebody of the +45 m sublevel is excavated the damage anglehas been deflected and the destruction and caving of thehanging wall rock mass gradually play a major role In theexcavation process of the first 60m the subsidence troughis basically located at the center of the mining area This isthe result of the interaction of sliding and toppling damageunder the influence of joints In subsequent excavations (80sim100 m) the damage development rate of the hanging wallis obviously higher than that of the footwall affected byjoints The subsidence trough is gradually deflected to thehanging wallThe comprehensive analysis has shown that thejoints and mining size are two important factors that causedifferent subsidence mechanisms at different mining stagesIt can also be observed that failure potentially occurs at thehangingwall zonewith continuousminingWith the decreasein the bearing capacity of the sidewall rock mass inducedby mining the subsidence range will be further expandedAfter the mining of the +30 m +15 m and 0 m sublevelsis finished the haulage road is bound to be destroyed andlarge-scale subsidence will occur at the surface of the hangingwall

7 Conclusions

In situ geological surveys and numerical analyses wereconducted to research the mechanism of rock mass caving


and surface subsidence in the Xiaowanggou iron mine Themain conclusions were the following(1) Based on an in situ geological survey the entire caving

project can be described as alternating development withslow sudden caving and the roof strata activity in the latermonitoring period is more intense With the continuousextraction the subsidence center slowly migrated to thehanging wall and the maximum subsidence rate decreasedThemaximum horizontal displacement was -2857mm in thehanging wall and the horizontal deformation of the hangingwall was significantly higher than that of the footwall Themining of the +30 m and +15 m sublevel ore bodies willincrease the risk of road breakage and collapse(2) Based on the numerical simulation the crack failure

occurs mainly along the randomly distributed joints andextends into the intact rock bridges when one or more keyrock bridges are damaged the stress arch will break immedi-ately When the rock mass is not completely collapsed to thesurface the damage angle of the hanging wall is always largerthan that of the footwall After the rock mass collapse to thesurface the damage development rate of the hanging wallis obviously higher than that of the footwall Ultimately thesubsidence trough is gradually deflected to the hanging walland large-scale subsidence will occur at the surface of thehanging wall(3) The arched-caving development of the rock mass is

verified in the study the root cause of the rock mass cavingis that the shear and tensile cracks are interlinked alongjoints and intact rock bridges The intrinsic factor of the slowsudden caving being periodic is the result of the stress archevolutionThe rock mass will not collapse to the surface untilthe last stress arch breaks(4) The development of surface subsidence is seriously

affected by joints and mining size The numerical result isbasically consistent with the result of the on-site monitoringit is feasible and reasonable to use the damage angle to analyzethe characteristics of surface subsidence With the continuedmining of the ore bodies the subsidence area will be furtherexpanded and the haulage road on the footwall is bound to bedestroyed It is necessary to take protective measures as soonas possible(5) Combined with the in situ geological investigation

and numerical simulation the rock mass caving mechanismand ground subsidence characteristics in a jointed rock massare clarified This research can be significant for miningsafety allowing mining technicians to better understand theinternal causes of rock caving and surface subsidence this canprevent the hazards caused by the caving and subsidence inadvance

Data Availability

The data used to support the findings of this study areavailable from the corresponding author upon request

Conflicts of Interest

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

Authorsrsquo Contributions

Fengyu Ren Dongjie Zhang andMiao Yu contributed to theformulation of overarching research goals and aims DongjieZhang and Jianli Cao analyzed the data Dongjie Zhangwrotethe paper Dongjie Zhang and Shaohua Li contributed to thenumerical calculation and simulation

Acknowledgments

The study is jointly supported by grants from the KeyProgram of National Natural Science Foundation of China(Grant no 51534003) National Key Research and Devel-opment Program of China (Grant no 2016YFC0801601)National Key Research and Development Program of China(Grant no 2016YFC0801606) The authors are grateful forthese supports

References

[1] V Lapcevic and S Torbica ldquoNumerical investigation of cavedrock mass friction and fragmentation change influence ongravity flow formation in sublevel cavingrdquo Minerals vol 7 no4 p 56 2017

[2] Y Abolfazlzadeh and M Hudyma ldquoIdentifying and Describinga Seismogenic Zone in a Sublevel CavingMinerdquoRockMechanicsand Rock Engineering vol 49 no 9 pp 1ndash17 2016

[3] J Blachowski and S Ellefmo ldquoNumerical modelling of rockmass deformation in sublevel caving mining systemrdquo ActaGeodynamica et Geomaterialia vol 9 no 3 pp 379ndash388 2012

[4] S Cao W Song D Deng Y Lei and J Lan ldquoNumerical sim-ulation of land subsidence and verification of its character foran iron mine using sublevel cavingrdquo International Journal ofMining Science and Technology vol 26 no 2 pp 327ndash332 2016

[5] M Cai P K Kaiser and C D Martin ldquoQuantification of rockmass damage in underground excavations from microseismicevent monitoringrdquo International Journal of Rock Mechanics andMining Sciences vol 38 no 8 pp 1135ndash1145 2001

[6] G Hu T Yang J Zhou et al ldquoMechanism of surrounding rockfailure and crack evolution rules in branched pillar recoveryrdquoMinerals vol 7 no 6 2017

[7] E Hoek and E T Brown Underground Excavation in RockInstitution of Mining and Metallurgy London UK 1980

[8] A R Bye and F G Bell ldquoStability assessment and slope design atSandsloot open pitSouth Africardquo International Journal of RockMechanics andMining Sciences vol 38 no 3 pp 449ndash466 2001

[9] H Li and R Brummer ldquoAnalysis of pit wall failure mechanismand assessment of long-term stability of pit walls Palaboraminerdquo Itasca Consulting Canada Ltd Technical report Min-nesota MN USA 2005

[10] K-S Woo E Eberhardt D Elmo and D Stead ldquoEmpiricalinvestigation and characterization of surface subsidence relatedto block cave miningrdquo International Journal of Rock Mechanicsand Mining Sciences vol 61 no 61 pp 31ndash42 2013

[11] S S Peng W M Ma and W L Zhong Surface subsidenceengineering 1992

[12] P-L Huang and C-X Chen ldquoAnalysis of mechanism of surfacesubsidence caused by underground mining with thick overbur-denrdquo Rock and Soil Mechanics vol 31 no 1 pp 357ndash362 2010


[13] E T Brown ldquoBlock caving geomechanicsrdquo inThe InternationalCaving Study Stage1 1997ndash2001 Julius Kruttschnitt MineralResearch Centre University of Queensland 2003

[14] H Zhao F Ma J Xu and J Guo ldquoPreliminary quantitativestudy of fault reactivation induced by open-pit miningrdquo Inter-national Journal of Rock Mechanics andMining Sciences vol 59pp 120ndash127 2013

[15] D Beck S Arndt I Thin C Stone and R Butcher ldquoA con-ceptual sequence for a block cave in an extreme stress anddeformation environmentrdquo in Proceedings of third internationalseminar on deep and high stress mining p 16 Quebec City 2006

[16] AVyazmenskyD Elmo andD Stead ldquoRole of rockmass fabricand faulting in the development of block caving induced surfacesubsidencerdquo Rock Mechanics and Rock Engineering vol 43 no5 pp 533ndash556 2010

[17] J Ju and J Xu ldquoSurface stepped subsidence related to top-coalcaving longwall mining of extremely thick coal seam undershallow coverrdquo International Journal of Rock Mechanics andMining Sciences vol 78 pp 27ndash35 2015

[18] Q Sun J Zhang Q Zhang and X Zhao ldquoAnalysis and pre-vention of geo-environmental hazards with high-intensive coalmining A case study in Chinarsquos western eco-environmentfrangible areardquo Energies vol 10 no 6 p 786 2017

[19] Y Huang J Zhang W Yin and Q Sun ldquoAnalysis of OverlyingStrataMovement and Behaviors in Caving and Solid BackfillingMixed Coal Miningrdquo Energies vol 10 no 7 p 1057 2017

[20] K Xia C Chen H Fu et al ldquoDeformation analysis of rockmass at different mining levels in west area of Chengchao ironrdquoJournal of Rock Mechanics and Engineering vol 35 no 4 pp792ndash805 2016

[21] J-A Wang J Tang and S-H Jiao ldquoSeepage prevention ofmining-disturbed riverbedrdquo International Journal of RockMechanics and Mining Sciences vol 75 pp 1ndash14 2015

[22] Y Zhao T Yang M Bohnhoff et al ldquoStudy of the rock massfailure process and mechanisms during the transformationfrom open-pit to underground mining based on microseismicmonitoringrdquo Rock Mechanics and Rock Engineering vol 51 no5 pp 1ndash21 2018

[23] H Zhao F Ma J Xu and J Guo ldquoIn situ stress field inversionand its application in mining-induced rock mass movementrdquoInternational Journal of Rock Mechanics and Mining Sciencesvol 53 pp 120ndash128 2012

[24] N Xu J Zhang H Tian G Mei and Q Ge ldquoDiscrete ele-ment modeling of strata and surface movement induced bymining under open-pit final sloperdquo International Journal of RockMechanics and Mining Sciences vol 88 pp 61ndash76 2016

[25] L He and Q B Zhang ldquoNumerical investigation of archingmechanism to underground excavation in jointed rock massrdquoTunnelling and Underground Space Technology vol 50 pp 249ndash261 2015

[26] T Szwedzicki ldquoRock mass behaviour prior to failurerdquo Interna-tional Journal of Rock Mechanics and Mining Sciences vol 40no 4 pp 573ndash584 2003

[27] D Elmo A Vyazmensky D Stead and J R Ranee ldquoA hybridFEMDEM approach to model the interaction between open-pit and underground block-caving miningrdquo in Proceedings ofthe 1st Canada-US RockMechanics Symposium - RockMechanicsMeeting Societyrsquos Challenges and Demands pp 1287ndash1294 Van-couverCanada May 2007

[28] C A Tang ldquoNumerical simulation of progressive rock fail-ure and associated seismicityrdquo International Journal of RockMechanics and Mining Sciences vol 34 pp 249ndash261 1997

[29] ldquoRFPA User Manualrdquo Dalian Mechsoft Co Ltd Dalian China2005

[30] L Lianchong Y Tianhong L Zhengzhao Z Wancheng andT Chunan ldquoNumerical investigation of groundwater outburstsnear faults in underground coal minesrdquo International Journal ofCoal Geology vol 85 no 3-4 pp 276ndash288 2011

[31] L C Li C A Tang W C Zhu and Z Z Liang ldquoNumericalanalysis of slope stability based on the gravity increasemethodrdquoComputers amp Geosciences vol 36 no 7 pp 1246ndash1258 2009

[32] N Li S Chang and F Y Ren ldquoHigh efficient divisional miningtechnology during transition from open-pit to undergroundmining in Xiaowanggou Iron Minerdquo Mental Mine vol 43 no5 pp 21ndash23 2014

[33] P Wang T Yang T Xu M Cai and C Li ldquoNumerical analysison scale effect of elasticity strength and failure patterns ofjointed rockmassesrdquoGeosciences Journal vol 20 no 4 pp 539ndash549 2016

[34] X Li S J Wang T Y Liu and F S Ma ldquoGround surfacemovement and fissures induced by underground mining in theJinchuan Nickel Minerdquo Engineering Geology vol 76 pp 93ndash1072004

[35] Z Fang and J P Harrison ldquoDevelopment of a local degradationapproach to the modelling of brittle fracture in heterogeneousrocksrdquo International Journal of Rock Mechanics and MiningSciences vol 39 no 4 pp 443ndash457 2002

[36] G W Ma X J Wang and F Ren ldquoNumerical simulation ofcompressive failure of heterogeneous rock-like materials usingSPH methodrdquo International Journal of Rock Mechanics andMining Sciences vol 48 no 3 pp 353ndash363 2011

[37] W Weibull ldquoA statistical distribution function of wide applica-bilityrdquo Journal of Applied Mechanics vol 18 no 3 pp 293ndash2971951

[38] C Jia Y Li M Lian and X Zhou ldquoJointed surroundingrock mass stability analysis on an underground cavern in ahydropower station based on the extended key block theoryrdquoEnergies vol 10 no 4 p 563 2017

[39] K S Kalenchuk S McKinnon and M S Diederichs ldquoBlockgeometry and rockmass characterization for prediction ofdilution potential into sub-level cave mine voidsrdquo InternationalJournal of Rock Mechanics and Mining Sciences vol 45 no 6pp 929ndash940 2008

[40] C Wang D D Tannant and P A Lilly ldquoNumerical analysisof the stability of heavily jointed rock slopes using PFC2DrdquoInternational Journal of Rock Mechanics and Mining Sciencesvol 40 no 3 pp 415ndash424 2003

[41] M A Mahtab D D Bolstad and F S Kendorski ldquoAnalysis ofthe geometry of fractures in San Manuel copper minerdquo Bureauof Mines Arizona AZ USA 1973

[42] NMohammad D Reddish and L Stace ldquoThe relation betweenin situ and laboratory rock properties used in numericalmodellingrdquo International Journal of RockMechanics andMiningSciences vol 34 no 2 pp 289ndash297 1997

[43] G A Nicholson and Z T Bieniawski ldquoA nonlinear deforma-tion modulus based on rock mass classificationrdquo InternationalJournal of Mining and Geological Engineering vol 8 no 3 pp181ndash202 1990

[44] Z T Bieniawski ldquoEngineering classification of jointed rockmassesrdquo Civil Engineering = Siviele Ingenieurswese vol 15 no12 pp 335ndash343 1973


[45] Z Bieniawski ldquoDetermining rock mass deformability experi-ence from case historiesrdquo International Journal of RockMechan-ics andMining Sciences amp Geomechanics Abstracts vol 15 no 5pp 237ndash247 1978

[46] L C Li C A Tang Z Z Liang T H Ma and Y B ZhangldquoNumerical simulation on water inrush process due to activa-tion of collapse columns in coal seam floorrdquo Journal of Miningand Safety Engineering vol 26 pp 158ndash162 2009 (Chinese)

[47] T-F Wong R H C Wong K T Chau and C A TangldquoMicrocrack statistics Weibull distribution and micromechan-ical modeling of compressive failure in rockrdquo Mechanics ofMaterials vol 38 no 7 pp 664ndash681 2006

[48] L C Li C A Tang and M Cai ldquoNumerical study on cavingstate of overlying rock and orebody by caving methodrdquoMentalMine vol 40 no 12 pp 13ndash17 2011

[49] L M Fan and R Q Huang ldquoProbability model for estimatingconnectivity rate of discontinuities and its engineering applica-tionrdquo Chinese Journal of Rock Mechanics and Engineering vol22 no 5 p 723 2003 (Chinese)

[50] G Meng D Fang L Li K Huang and Y Jiang ldquoStudyof equivalent strength parameters of ubiquitous joint modelfor engineering rock mass with preferred intermittent jointsrdquoChinese Journal of Rock Mechanics and Engineering vol 32 no10 pp 2115ndash2121 2013 (Chinese)

[51] R X He F Y Ren D L Song X You and Y Fu ldquoInducedcaving rule of inclined thick ore body in Hemushan iron minerdquoJournal of Mining and Safety Engineering vol 34 pp 899ndash9042017 (Chinese)

[52] C C Li ldquoRock support design based on the concept of pressurearchrdquo International Journal of Rock Mechanics and MiningSciences vol 43 no 7 pp 1083ndash1090 2006

[53] M Li H Zhang W Xing Z Hou and P Were ldquoStudy of therelationship between surface subsidence and internal pressurein salt cavernsrdquo Environmental Earth Sciences vol 73 no 11 pp6899ndash6910 2015

[54] T Villegas E Nordlund and C Dahner-Lindqvist ldquoHanging-wall surface subsidence at the Kiirunavaara mine SwedenrdquoEngineering Geology vol 121 no 1-2 pp 18ndash27 2011

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[6] and thus induce massive caving of the roof strata Theformation of surface subsidence is the inevitable result of thecaving to the surface

The rock mass failure and surface subsidence caused byundergroundmining depend onmany factors such asminingdepth geological conditions and rock mass structure [14ndash16] Many studies have conducted in-depth analysis of themining-induced rock mass failure and surface subsidencewhich can be divided into three categories The first categoryis the physical similarity simulation method By establishinga model similar to the actual engineering the displacementstress and failure under the influence of mining can beanalyzed Ju et al [17] studied the surface stepped subsidencemechanism and proposed the important role of primarykey stratum and subkey stratum in the development ofsubsidence Sun et al [18] studied the characteristics ofroof strata movement and the evolution of water-conductingfractures and proposed SBM technology to effectively con-trol strata movement and deformation Huang et al [19]studied the characteristics of the goaf-roof displacement andstrain through physical similarity experiments and reason-ably determined the key parameters required for miningdesign The second category is on-site monitoring methodsuch as total station surveys GPS field surveys differentialInSAR techniques borehole monitoring and microseismicmonitoring Xia et al [20] used borehole monitoring tech-nology to analyze the caving process of the goaf-roof andpointed out that the caving evolution was affected by thestructural character of the rock mass Wang et al [21] used acombination of on-sitemonitoring and numerical simulationto analyze the fracture development of the rock mass andsurface subsidence characteristics near the riverbed andproposed the Four-Element Structure of antiseepage for theriverbed Zhao et al [22] used microseismic technology toanalyze the stability of goafs and slope during the transitionfrom open-pit to underground mining and to study the rockfailure mechanism The third category is numerical simula-tionmethods such as the finite elementmethod (FEM) finitedifference method (FDM) and discrete element method(DEM) which have developed rapidly in recent years Zhao etal [23] analyzed the rock movement mechanism and surfacesubsidence characteristics under different stress fields Xuet al [24] proposed an EJRM-based numerical method tostudy the strata and surface movement under the influenceof a jointed rock mass He et al [25] used a discontinuousdeformation analysis to study the formation and stabilizationof a stress arch in laminated rock mass

Based on the above analysis results each method has itsown advantages However the physical similarity simulationmethod is mainly focused on underground coal mining andthe rock structure of metal mines is different from that ofcoal mines [23] The on-site monitoring method is mainlyused to evaluate and analyze the trend of rock caving andsurface subsidence but cracks and stress evolution in therock mass cannot be represented visually When we usenumerical methods for analysis the simulation results maynot be consistent with actual site conditions [26 27] Theneeds of accurate study of the mechanism of caving andsubsidence cannot bemet solely using onemethod and there

are relatively few studies on the cavingmechanism of the rockmass in metal mines Therefore using a combination of on-site monitoring and numerical simulation it is possible torealize a comprehensive explanation of the sublevel cavingmining from the macroscopic rock mass collapse modeland surface subsidence trend of the caving mechanism ofroof strata and the microscopic crack propagation law Thisconnects mining and rock failure in time and space thusproviding support for mining safety and the protection ofsurface environment and facilities

In this paper the Xiaowanggou iron mine was selected asan engineering project for an in-depth study The rock masscaving mechanism and surface subsidence characteristicsduring the sublevel caving mining were comprehensivelystudied using an in situ geological investigation and numer-ical analysis method (the FEM-based numerical softwareRFPA 2D [28ndash31]) The overall understanding of the processfrom the crack propagation stress evolution and damagedevelopment to the rock caving and surface subsidence isobtained The research can provide guidance for other metalmines with similar conditions

2 Background of Geologic andMining Conditions

The Xiaowanggou iron mine is a sedimentary metamorphicmagnetite deposit and is located in Liaoyang City LiaoningProvince northeast ChinaThe mining area is approximately70 km long and the elevation is approximately 250m It hasa temperate continental monsoon climate with an averageannual precipitation of approximately 800mmThe ore bodyis stratoid and the strike is north by east 50∘ with the dipdirection of south-east and the dip angle ranging from 30∘ to50∘ The strike length is approximately 350m and the averagethickness is 100m Thus it is a typically inclined metal minethe deep mining area extends from 190 m below the surfaceto over 340 m and the surrounding rock is mainly mixedgranite

The Xiaowanggou iron mine is mined by partition Theentire deposit from top to bottom is divided into threeminingareas namely the uppermining area themiddlemining areaand the deep mining area and there is a safety pillar betweenthe partitions [32] The deep mining area was selected as theresearch object of this studyThe sublevel caving method wasused in this mining area and the first mining section waslocated at the +60m level with a subsection height of 15mAt present the major mining activities are carried out at the+30m level and the exploration system has been arrangedto the 0 m level In the mining process the overlying strataface continuous fracturing and caving In the late stage of+45m sublevel mining an oval collapse pit with a maximumdepth of 27m and a diameter of about 80mwas formed on theground in October 2015 Multiple fracture lines were formedaround the collapse pit which poses a serious threat to theenvironment surface industrial facilities and haulage roads(Figure 1)

This is of great significance for grasping the process ofrock mass caving and the trend of surface subsidence inadvance which can effectively reduce the potential safety


ChinaBeijing

Liaoyang

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

4

2480

2423

2550

2574

1

2 3

4

50m

C1

C3

C2

C6

C4

C5



Ore body boundary

Observation cracks

Monitoring station

Monitoring borehole

Haulage road

Exploration line

05m

Collapse pit

27m

Cracks

80m

Road

Collapse pit

Crack

Cracks

North









1 2391 891 1502 2325 825 1503 2341 841 1504 2363 863 150

1st mining layer

Monitoring borehole

Ore body boundary

2nd mining layer

3rd mining layer

Mixed granite

2150 2250 2350 24500m

60m

100m

200m

250mE W

247m

0m

100m

45m

15m30m

1 2 3 4













53 m


Fracture

(a)

126 m


Fracture

(b)


159

25

159

20

159

24

159

14

156

18

155

15

152

17

148

15

151

16



1105

624

589

353245

93

481m

236m

152Suddenly caving

Slowly caving

0

20

40

60

80

100

120

Roof

rock

cavi

ng h

eigh

t (m

)

(a)


140815140916141013141115141217150116150217

150315150417150515150618150716150815150914150920150924150925

Caving arch


Initial caving line

Late caving line

0

20

40

60

80

100

120

140

Roof

rock

cavi

ng h

eigh

t (m

)

(b)



2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325



Mining direction

North

4Haulage road









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









ChinaBeijing

Liaoyang

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

4

2480

2423

2550

2574

1

2 3

4

50m

C1

C3

C2

C6

C4

C5



Ore body boundary

Observation cracks

Monitoring station

Monitoring borehole

Haulage road

Exploration line

05m

Collapse pit

27m

Cracks

80m

Road

Collapse pit

Crack

Cracks

North









1 2391 891 1502 2325 825 1503 2341 841 1504 2363 863 150

1st mining layer

Monitoring borehole

Ore body boundary

2nd mining layer

3rd mining layer

Mixed granite

2150 2250 2350 24500m

60m

100m

200m

250mE W

247m

0m

100m

45m

15m30m

1 2 3 4













53 m


Fracture

(a)

126 m


Fracture

(b)


159

25

159

20

159

24

159

14

156

18

155

15

152

17

148

15

151

16



1105

624

589

353245

93

481m

236m

152Suddenly caving

Slowly caving

0

20

40

60

80

100

120

Roof

rock

cavi

ng h

eigh

t (m

)

(a)


140815140916141013141115141217150116150217

150315150417150515150618150716150815150914150920150924150925

Caving arch


Initial caving line

Late caving line

0

20

40

60

80

100

120

140

Roof

rock

cavi

ng h

eigh

t (m

)

(b)



2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325



Mining direction

North

4Haulage road









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018











1 2391 891 1502 2325 825 1503 2341 841 1504 2363 863 150

1st mining layer

Monitoring borehole

Ore body boundary

2nd mining layer

3rd mining layer

Mixed granite

2150 2250 2350 24500m

60m

100m

200m

250mE W

247m

0m

100m

45m

15m30m

1 2 3 4













53 m


Fracture

(a)

126 m


Fracture

(b)


159

25

159

20

159

24

159

14

156

18

155

15

152

17

148

15

151

16



1105

624

589

353245

93

481m

236m

152Suddenly caving

Slowly caving

0

20

40

60

80

100

120

Roof

rock

cavi

ng h

eigh

t (m

)

(a)


140815140916141013141115141217150116150217

150315150417150515150618150716150815150914150920150924150925

Caving arch


Initial caving line

Late caving line

0

20

40

60

80

100

120

140

Roof

rock

cavi

ng h

eigh

t (m

)

(b)



2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325



Mining direction

North

4Haulage road









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









53 m


Fracture

(a)

126 m


Fracture

(b)


159

25

159

20

159

24

159

14

156

18

155

15

152

17

148

15

151

16



1105

624

589

353245

93

481m

236m

152Suddenly caving

Slowly caving

0

20

40

60

80

100

120

Roof

rock

cavi

ng h

eigh

t (m

)

(a)


140815140916141013141115141217150116150217

150315150417150515150618150716150815150914150920150924150925

Caving arch


Initial caving line

Late caving line

0

20

40

60

80

100

120

140

Roof

rock

cavi

ng h

eigh

t (m

)

(b)



2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325



Mining direction

North

4Haulage road









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









2480

2550

2574

C1

C3

C2

C6

C4

C5N1 N2

N3N4

N5N6 N7

N8N9

N10N11

N12N13

N14

R1R2

R3R4

R5R6

R7

R8R9R10

R11R12

R13R14

R152015523

40004000

4400 4400

2000

2000

2500

2500

2880

2648

26231

2

3

2014727

2014928

20141130

2015128

2015325



Mining direction

North

4Haulage road









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









Distance (m)

280240120 20080 16040

140622140823141022141220

150119150322150521150720




0

Surf

ace s

ubsi

denc

e (m

m)

(a)

28024020016012040 80

114m

+45m miningarea

Haulage road


minus500

50100150200

Hor

izon

tal d

ispl

acem

ent (

mm

)

Distance (m)

140622140823141022141220

150119150322150521150720

(b)

28024020016012040 80

Hor

izon

tal d

efor

mat

ion

(mm

m)

+45m mining area

Haulage road

140622140823141022141220

150119150322150521150720


0123456789

10

Distance (m)

(c)

N8

Mining direction

0

2

4

6

8

10

Subs

iden

ce ra

te (m

md

)


0

200

400

600

800

1000

Tota

l sub

side

nce (

mm

)


(d)



119863119909 = 1198630 [1 minus exp(minus119898 (119909 + 119897)V)]119899

(1)



V119909 = 1198631199091015840

= 1198981198991198630 exp(minus119898 (119909 + 119897)V)


(2)



Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









Surface cracks



4 Modeling Work



120601 = 1198981205830 (1205831205830)119898minus1

exp [minus( 1205831205830)119898

] (3)



119864 = (1 minus 119863) sdot 1198640 (4)


tu t0

minusft0

minusftr

(a)

fc0

fcr

c0

(b)



119863 =

0 120576 gt 12057611990501 minus 1205821205761199050120576 1205761199050 ge 120576 gt 1205761199051199061 120576 le 120576119905119906

(5)








1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018











1 294 70 163 041 1072 225 5 095 027 0653 043 81 187 035 122





119863 =

0 120576 lt 12057611988801 minus 120582 sdot 1205761198880120576 120576 ge 1205761198880

(7)







119864 = 10(119877119872119877minus10)40 (8)






Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018











Rock type




PoissonrsquosratioV



Mixedgranite 35 553 024 30 2700



1198911198881198911198880 = 02602 ln119898 + 00233 12 le 119898 le 50 (9)

1198641198640 = 01412 ln119898 + 06476 12 le 119898 le 10 (10)








Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018










Rock type


Homogeneityindex m




PoissonrsquosratioV



Mixedgranite 3 11 69 69 024 30 2650

Joints 5 1 10 1 032 30 1000

549 MPa

124 MPa

5 10 15 20 25 30 350Load steps (n)

0

10

20

30

40

50

60

70

Com

pres

sive s

tren

gth

(MPa

)








250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









250m

O X

Z

15m

Undercut

Joints

250m 250m

175m

Ore body


600m










Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









Cracks initiation


72∘ 76∘

(a)



70∘ 73∘

(b)


Sliding failure

68∘ 71∘

(c)


Stress arch

Caving arch

66∘ 67∘

(d)




64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018










64∘62∘

(a)


59∘61∘

(b)


59∘54∘

(c)


56∘52∘

(d)



Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018









Distance (m)




-5824cm

minus1000

minus800

minus600

minus400

minus200

0

Surf

ace s

ubsi

denc

e (cm

)

0 40 80 120 160 200 240 280 320

+45m mining area






Inflection point


the hanging wall

50

55

60

65

70

75

80

Dam

age a

ngle

s (D

egre

e)





7 Conclusions









Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018















Data Availability






Acknowledgments


References
































































Journal of












Journal of







Volume 2018






Volume 2018



























































Journal of












Journal of







Volume 2018






Volume 2018








study on the rock mass caving and surface subsidence...

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