Support System Of Bord And Pillar Workings Engineering Essay

Support System Of Bord And Pillar Workings Engineering Essay

The appraisal of stone burden from the strata and its distribution over the belowground mine workings is of premier importance. In Indian coalmines, CMRI- RMR and NGI-Q Systems are largely used for explicating design of support in stone technology. In this undertaking work their appraisal is been done and design of support system done by CMRI – RMR & A ; Q system.

The design of support system besides done with the aid of numerical mold by imitating the workings. Study and analysis of the emphasis distribution around development and depillaring workings in coal mines and perpendicular emphasis appraisal is done.

The working has been modelled by composing a plan codification in FLAC5.0. The modeling is done for merely drive of galleries ( development ) to organize three pillars and support system designed for them. The perpendicular and emphasis contours besides plotted.

The plotting stess show that the ultimate perpendicular emphasis additions well with addition in the deepness screen and acquire concentrated over the country of digging with high emphasis concentration over the pillars, stooks and ribs above the normal emphasis under the given deepness screen.

Chapter 1

Introduction

Aim

Aim of the present work is to plan support system in development and depillaring workings by conventional method and numerical mold of support system for development workings.

Design parametric quantities

Strata behaviour

Depth of screen

Method of extraction

Equipment choice for digging

Span of country

Height of extraction

To plan support systematic support regulations ( SSR ) should be followed. Planing optimized support system means we have given proper support non less or non more, it helps in cost control of supports.

1.1 PRESENT SUPPORT SYSTEM

Goaf borders

At goaf borders cogs shall be set tegument to clamber. Prop up shall be set in between cogs, cogs & A ; coal sides.

Working faces

At working faces props shall be set at a maximal interval of 1.2 m between the rows of props or in the same row.

Cogs shall be set at all entrywaies to the countries under extraction & A ; besides at interval of non more than 2.4m in the country under existent extraction.

Areas near to the faces where supports are likely to be affected due to blaring shall be supported by cross bars.

Support of galleries

Props shall be set at interval of 1.2 m between them in the same row & A ; at a soap interval between rows of props in all galleries & A ; splits with in a distance of 2 pillars from the pillar under extraction or a distance of 30m, which of all time greater.

Cogs shall be set at all junctions.

Wider gallery

Gallery & gt ; 4.8 metre shall be supported with cogs at interval non transcending 2.4m between cogs & A ; between rows of cogs.

1.2 ACCIDENT STATISTICS OF INDIAN UNDERGROUND COAL MINES

India has big resources of coal sedimentations for resistance excavation & A ; batch of coal was blocked in bing belowground mines. Safe extraction of these can be made possible by effectual strata direction & A ; proper support design. Accident due to motion of strata in belowground coal mines had been a major concern for the excavation industry & A ; its largest lending factor of belowground coal mine accident. Continuous effeorts were being made by all concerns to cut down jeopardy of strata motion. To cut down strata motion monitoring of strata & A ; proper design of support system is indispensable.

Year

Fall of roof

Fall of sides

Entire

Entire B/G ACC

Percentage of accidents due to strata motion

1997

38

12

50

94

53

1998

35

15

50

80

62

1999

33

11

44

74

59.5

2000

27

14

41

62

66

2001

30

9

39

67

58

2002

23

11

34

48

70

2003

18

5

23

46

50

2004

26

8

34

49

69

2005

18

7

25

49

51

2006

13

4

17

44

40

2007

13

4

17

25

68

2008

13

7

20

33

60

Entire

287

107

394

671

59.1

Table 1: Cause wise fatal accidents in coal mines, due to strata motion

The analysis of strata due to strata motion for last 12 old ages ( 1997- 2008 ) revealed that:

The roof autumn & A ; side autumn accidents accounted for 59 % of all below land fatal accidents in coal mines.

Accidents due to fall of roof occurred in about same proportion in bord & A ; pillar development every bit good as depillaring territories.

The cause of the roof autumn is due to improper design of support system. So in order to diminish the accident & A ; increase productiveness we need to plan proper support system.

Chapter 2

LITERATURE REVIEW

2.1 Rock burden

Maximal burden ( P ) that is required to be supported in the split and piece can be estimated utilizing the undermentioned expression and as detailed elsewhere [ Kushwaha, 2005 ] :

P = I? . SF1.5h aˆ¦aˆ¦aˆ¦aˆ¦aˆ¦aˆ¦.. ( 8 )

Where, I? = leaden norm stone denseness, 2.5 t/m3 ( carbonous shale )

SF1.5h = tallness of safety factor contour up to 1.5 in the roof strata in the fake theoretical account.

2.2 SUPPORT ESTIMATION

A form of support may be proposed utilizing the following expression such that an equal support safety factor ( about 1.1-1.25 in depillaring countries, about 1.5-2.0 for lasting roadways ) is achieved:

Where, n = the figure of bolts/props in a row

bc= to the full column grouted roof bolt capacity, 8 metric ton to the full column rosin roof bolt

capacity, 16 metric ton capacity of lumber props, 10 metric ton capacity of timber cogs, 20

metric ton

w =width of the piece, here 4.2m

sp= spacing between two rows

Support safety factor = S / P

2.3 GUIDELINES FOR DRAWING OF SUPPORT PLANS IN BORD & A ; PILLAR

WORKINGS IN COAL MINES

General:

The assorted phases of planing a suited support system and guaranting successful installing are fundamentally as follows:

( a ) A geotechnical study and reading of study findings

( B ) Selection/designing of support system based on above reading

( degree Celsius ) Choice of equipment

( vitamin D ) Actual installing procedure and

( vitamin E ) Monitoring of the system.

Two systems are peculiarly used to qualify excavation land conditions.

2.31 BARTON ‘S Q-SYSTEM ( Rock quality index, Norse Geotechnical Institute )

It is evaluated as

RQDx Junior x jw

Q = — — — — — — — — — –

jn X ja X SRF

Where RQD = stone quality appellation

Jn = joint set figure

Junior = joint raggedness figure

ja = joint change figure

jw = joint H2O decrease figure

SRF = emphasis decrease factor.

Based on the value of Q the stone mass can be described as “ exceptionally good ” ( Q=400 to 1000 ) to “ exceptionally hapless ” ( Q=0.001 to 0.01 ) . Using the Q value, the upper limit unsupported span of roof can be estimated by the expression:

Span ( m ) = 2 ten ESR x Q 0.4

Where ESR is excavation support ratio ( which is 3 to 5 for impermanent mine workings and 1.6 for lasting workings ) . The stone burden ( Proof ) can be estimated from the empirical expression:

2.0 x F

Proof ( t/m2 ) = — — — — — — —

Jr x Q 0.33

Where F = 1 if Jr is 9 or more

Or F= ( ( Jr 0.5 ) /3 if Jr is less than 9.

Depending on the different values of the parametric quantities and Q, 38 support classs have been identified.

2.32 BIENIAWSKI ‘S RMR SYSTEM

There are five parametric quantities in this categorization:

( I ) Intact stone strength

( two ) RQD

( three ) Joint spacing

( four ) Condition of articulations

( V ) Ground H2O ooze

Rating division for each of the parametric quantities is given and RMR is amount of five evaluations. Based on

RMR, the roof is classified as really good ( RMR:80-100 ) to really hapless ( RMR:0-20 ) . From this appraisal of stone burden is derived utilizing theoretical relation and support usher is provided.

Suitability of Q-system/ RMR system

These two categorizations have been applied to about 30 Indian coal mines. The Q categorization is suited for extremely jointed stones for difficult stone conditions. Most of the parametric quantities in this system are based on joint properties whereas stableness in coal mines is non simply joint controlled. The SRF has no relation with the emphasis field happening about multiple gaps like coal mine roadways. The parametric quantity descriptions in Q system leave much to subjective judgements.

The RMR system gives consequences nearer to existent roof conditions. it was recognized that in the most of the Indian coal mines, bedding planes, structural characteristics and weathering of roof stones are so major causes of roof failure. In Bieniawski ‘s attack, consideration is non given to sedimentary characteristics, structural features other than articulations and weatherability of stones. Deviations in the consequences besides arise from the weightages for the parametric quantities which need to be adjusted to Indian stone conditions.

2.4 CMRI-RMR ROCK MASS CLASSIFICATION

This stone mass categorization system is being used on a regular basis by academic and research institutes.

The five parametric quantities used in the categorization system and their comparative evaluations are summarized below:

Sl no

Parameter

Max. evaluation

1

Layer thickness

30

2

Structural characteristics

25

3

Rock weatherability

20

4

Strength of roof stone

15

5

Land H2O ooze

10

Table 2. CMRI-RMR prescribed parametric quantities for RMR finding

The five parametric quantities should be determined separately for all the stone types in the roof upto a tallness of at least 2 m.

1. Puting thickness: Spacing between the bedclothes planes or planes of discontinuities should be measured utilizing borehole stratascope in a e m long drill made in the roof. Alternately, all bedding planes or weak planes within the roof strata can be measured in any roof exposure like a roof mistake country, shaft subdivision or cross step impetus. Core boring shall be attempted wherever executable and the nucleus log can be used to measure RQD and bed thickness. Average of five values should be taken and layer thickness should be expressed in centimeter.

2. Structural Features: Random geological function should be carried out and all the geological characteristics ( discontinuities like articulations, mistakes and faux pass, and sedimentary characteristics like cross bedclothes, sandstone channels ) should be carefully recorded. The comparative orientation, spacing and grade of copiousness for all these characteristics shall be noted. Their influence on gallery stableness should be assessed and the structural index for each characteristic should be determined from the Table 1 as given below.

3. Weatherability: ISRM standard slake lastingness trial should be conducted on fresh samples from the mine to find the susceptibleness of stones to enduring failure on contact with H2O or the atmospheric wet. For this trial, weigh precisely any 10 irregular pieces of the sample ( the sum weight should be between 450- 500 g ) ; put them in the trial membranophone immersed in H2O and revolve it for 10 min at 20 revolutions per minute ; dry the stuff retained in the membranophone after the trial and weigh it once more. Weight per centum of stuff remaining after trial is the concluding slake lastingness index, expressed in per centum. Mean of three such first rhythm values should be taken. Core may be broken to obtain the samples.

4. Rock Strength: Point burden trial is the standard index text for mensurating the strength of stones in the field. Irregular samples holding ratio of 2:1 for longer axis to shorter axis can be sued for the trial. The sample is kept between the pointed platens and the burden is applied gently but steadily. The burden at failure in kilogram divided by the square of the distance between the platens in centimeter gives the point burden index ( Is ) . The mean of the highest five values out of at least 10 sample trials should be taken. The compressive strength of the stones can be obtained from the irregular ball point burden index for Indian coal step stones by the relation:

Co = 14 Is ( in kg/cm2 )

5. Land H2O: A 2m long perpendicular borehole should be drilled in the immediate roof and the H2O oozing through the hole after half an hr should be collected in a measurement cylinder. The norm of three values from three different holes should be taken and expressed in ml/minute.

Rock Mass Rating ( RMR ) is the amount of five parametric quantity evaluations. If there are more than one stone type in the roof, RMR is evaluated individually for each stone type and the combined RMR is obtained as:

I? ( RMR of each bed x bed thickness )

Combined RMR = — — — — — — — — — — — — — — — — — — — — — — — —

I? ( Thickness of each bed )

The RMR so obtained may be adjusted if necessary to take history for some particular state of affairss

in the mine like deepness, emphasis, method of work

FIG.1 Flow-sheet for deducing RMR

Sl no

Rock mass evaluation

Rock quality

1

0 – 20

Very hapless

2

20 – 40

Poor

3

40 – 60

Carnival

4

60 – 80

Good

5

80 – 100

Very good

Table 3: Categorization of stone mass evaluation

Design OF SUPPORT FOR DEPILLARING WORKING

In general, Rock Mass Rating ( RMR ) is used for design of supports in development galleries. However, due to restrictions of its application to depillaring workings, many research workers adopted assorted attacks such as Q-classification of stone mass, numerical mold etc for design of support system in depillaring workings, Some times, it is besides required to plan support in a depillaring panel holding widely changing geo excavation conditions with different support denseness.

For the intent of support design in a typical depillaring country, Barton ‘s Rock mass categorization index- Q was besides determined as follows:

Q = { RQD/Jn } { Jr/Ja } { Jw/SRF } — — ( 1 )

Rock Quality Designation = degree Fahrenheit ( layer thickness ) = 97

Jn = no articulations were observed in the roof = 4 for galleries

= 12 for junctions

= 20 for goaf borders

Ja = Plant feelings are frequent in the roof ; nevertheless kettle bottoms/sandstone channels/slicken sides are non perceptible = 1

Jw = by and large dry up to 8 milliliter of H2O per minute seepage.= 1

Jr = Smooth planar articulations = 1

SRF values for assorted geometries during depillaring are as follows:

c/Ms SRF

For galleries and junctions: & gt ; 10 1

1 – 10 1-2

For pieces: & gt ; 5 2

2.5 – 5 3 – 5

& lt ; 2.5 5

For goaf borders: any value 10

Roof force per unit area could be estimated by the dealingss based on the Q value adjusted to the geometrical conditions:

For joint set figure ( Jn ) & gt ; 9, the roof force per unit area ( Proof ) = 2/Jr x ( 5Q ) -1/3

For Jn & lt ; 9, Proof = 2/3 Jn1/2 /Jr x ( 5Q ) -1/3

2.6 Monitoring AND CONTROL OF STRATA MOVEMENT:

CONVENTIONAL BORD AND PILLAR EXTRACTION

aˆ? To minimise the dangers from burdening on the pillar due to overhanging of roof in the goaf and to guarantee that every bit little an country of un-collapsed roof as possible is allowed in the goaf, a suited codification of pattern for induced blaring shall be evolved in audience with a scientific organisation maintaining in position the deepness of induce shooting holes being non less than 2.7 m, way & A ; spacing of shotholes, explosives used etc. so as to restrict the rate of convergence [ i.e. , the ratio of C1/C2 is equal or less than 2, where C1 is day-to-day convergence at a site in a twenty-four hours “ n ” and C2 is the mean day-to-day convergence at the site upto the old twenty-four hours i.e. twenty-four hours ( n-1 ) ] and besides to guarantee complete filling of the goaf and release of any abutment force per unit areas.

aˆ? Convergence recording Stationss shall be installed at all junctions situated within two pillar distance from pillar under extraction in the proposed panel. Monitoring of readings at convergence entering Stationss shall be done in every displacement by a competent individual punctually authorized by the director and the measurings shall be recorded in a edge paged book and the same shall be counter signed daily by the Under Manager of the displacement and Asst. Manager in charge. All the work individuals shall be withdrawn from the abutment zone if the ration of C1/C2 is equal to or more than 2 as given above and stairss shall be taken to let go of the goaf abutment force per unit area by induced blasting. The Safety Officer shall organize recording, analysis and reading of the readings and advises the Officers/ Officials daily at the mine.

FIG.2 TYPICAL INSTRUMENTS FOR STRATA Monitoring

NUMERICAL Modeling

A computing machine simulation, a computing machine theoretical account or a computational theoretical account is a computing machine plan, or web of computing machines, that attempts to imitate an abstract theoretical account of a peculiar system. Models can take many signifiers, including but non limited to dynamical systems, statistical theoretical accounts, differential equations, or game theoretic theoretical accounts.

Frequently when applied scientists analyze a system to be controlled or optimized, they use a mathematical theoretical account. In analysis, applied scientists can construct a descriptive theoretical account of the system as a hypothesis of how the system could work, or seek to gauge how an unforeseeable event could impact the system. Similarly, in control of a system, applied scientists can seek out different control attacks in simulations.

A mathematical theoretical account normally describes a system by a set of variables and a set of equations that set up relationships between the variables. The values of the variables can be practically anything ; existent or integer Numberss, Boolean values or strings, for illustration. The variables represent some belongingss of the system, for illustration, measured system end products frequently in the signifier of signals, clocking informations, contours, and event happening ( yes/no ) . The existent theoretical account is the set of maps that describe the dealingss between the different variables. Here FLAC ( Fast Lagrangian Analysis of Continua ) 5.0 has been used for simulation and analysis.

2.71FLAC 5.0

FLAC is a planar explicit finite difference plan for technology mechanics calculation. This plan simulates the behaviour of constructions built of dirt, stone or other stuffs that may undergo plastic flow when their output bounds are reached. Materials are represented by elements, or zones, which form a grid that is adjusted by the user to suit the form of the object to be modeled. Each component behaves harmonizing to a prescribed linear or nonlinear stress/strain jurisprudence in response to the applied forces or boundary restraints. The stuff can give and flux and the grid can deform ( in large-strain manner ) and travel with the stuff that is represented. The explicit, Lagrangian computation strategy and the mixed-discretization zoning technique used in FLAC guarantee that plastic prostration and flow are modeled really accurately. Because no matrices are formed, big planar computations can be made without inordinate memory demands. The drawbacks of the expressed preparation ( i.e. , little timestep restriction and the inquiry of needed damping ) are overcome to some extent by automatic inactiveness grading and automatic damping that do non act upon the manner of failure.

Though FLAC was originally developed for geotechnical and excavation applied scientists, the plan offers a broad scope of capablenesss to work out complex jobs in mechanics. Several constitutional constituent theoretical accounts that permit the simulation of extremely nonlinear, irreversible response representative of geologic, or similar, stuffs are available.

However, it offers several advantages when applied to technology jobs.

1. The input linguistic communication is based upon recognizable word commands that let you to place the application of each bid easy and in a logical manner ( e.g. , the APPLY bid applies boundary conditions to the theoretical account ) .

2. Engineering simulations normally consist of a drawn-out sequence of operations – e.g. , set up unmoved emphasis, use tonss, excavate tunnel, install support, and so on. A series of input bids ( from a file or from the keyboard ) corresponds closely with the physical sequence that it represents.

3. A FLAC informations file can easy be modified with a text editor. Several informations files can be linked to run a figure of FLAC analyses in sequence. This is ideal for executing parametric quantity sensitiveness surveies.

4. Theword-oriented input files provide an first-class agencies to maintain a documented record of the analyses performed for an technology survey. Often, it is convenient to include these files as an appendix to the technology study for the intent of quality confidence.

5. The command-driven construction allows you to develop pre- and post-processing plans to pull strings FLAC input/output as desired. For illustration, you may wish to compose a meshgeneration map to make a particular grid form for a series of FLAC simulations. This can readily be accomplished with the FISH scheduling linguistic communication, and incorporated straight in the input informations file

Chapter 3

Design OF SUPPORT SYSTEM

Development workings

depillaring territories

Table – : 4 Types of supports used in coal mines & A ; their burden bearing capacity

( CMRI Report, 1987 )

Sl no

Support point

Load bearing capacity ( T )

1

Roof bolt ( full column grouted with speedy puting cement capsules ) ( TMT ribbed bolt of 22 millimeters diameter )

6

2

Roof bolt ( full column grouted with rosin capsules ) ( TMT ribbed bolt of 22 millimeters diameter )

12

3

Roof sewing

8

4

Wooden prop

10

5

Steel prop

30

6

Steel wedge

30

7

Wooden wedge

20

8

Pit prop

15

DETERMINATION OF CMRI – RMR

Table 5: RMR computation of given informations

Sl No

Parameter

Description

Rating

1

Layer thickness

17cm

19

2

Structural characteristics

Joint faux pas ( indices = 8 )

14

3

Weatherability

91 %

11

4

Compressive strength

215 kg/cm2

05

5

Land H2O

Moist

09

5

Entire RMR

58 ( just stone )

DETERMINATION OF Q VALUE BY NGI – Q System

Table 5: computation of Q

SRF

Q

Parameters ( a )

RQD

Jn

Junior

Ja

Jw

pieces

Goaf borders

pieces

Value ( a )

60

9

1.5

1

1

5

10

2

3.1 GEOMINING Details

Seam thickness: 7 metres

Pillar size: 20m * 20m ( Centre to center )

Gallery size: 4.5 m * 3m

Depth of screen: 100 metres

RMR: 58 ( FAIR )

Slice breadth: 5m

3.2 Equations for support burden

Rock burden = B A- D*F ( 1.7-0.037 A- RMR + 0.0002 A- RMR ) 2aˆ¦aˆ¦aˆ¦aˆ¦ . ( 3.1 )

Where, B ( breadth of galleries/splits ) ,

D ( mean stone denseness )

F is safety factor

A safety factor of 1.5 is by and large considered plenty.

3.3 ESTIMATION OF ROCK LOAD & A ; DESIGN OF SUPPORT SYSTEM IN development working

Rock burden in galleries and splits ( BY CMRI-RMR )

Rock burden ( t/m2 ) in the galleries and splits in depillaring countries has been determined utilizing the empirical relationship of CMRI-RMR System i.e in ( equation 3.1 ) :

Rock burden = B A- D*F ( 1.7-0.037 A- RMR + 0.0002 A- RMR ) 2aˆ¦aˆ¦aˆ¦aˆ¦ . ( 3.1 )

RMR = 58

Gallery span ( B, breadth ) = 4.5 m

Density ( D ) = 2.29 t/m2

Height = 3m

Safety factor ( F ) = 2 ( by and large )

Rock burden = 4.80 t/m2

Hence, stone burden in galleries and splits = 4.80 t/m2

Design OF SUPPORT SYSTEM

To back up the stone burden we need to give support higher than the stone burden at that country so that the roof will non fall

It is assumed that full column rosin bolt is used as the support system.

We need to pick support from the tabular array 3.1

Roof bolt capacity = 12 T

Bolt spacing = 1.5 m

Distance between two rows of bolts = 1.5 m

No of bolts in a row= 3

The support design shown in the figure 3

( n * A ) + ( thousand * Q )

Applied support burden = — — — — — — — — — — — — — — aˆ¦aˆ¦aˆ¦aˆ¦aˆ¦aˆ¦ ( 3.2 )

W * a

Where, n is figure of bolts in a row

A is the anchorage strength of each bolt ( T )

Q is the burden bearing capacity of the extra support if done ( T )

m is the figure of extra support at spacing “ a ” if it has been used

W breadth of split or piece

a is the spacing between two back-to-back rows

The above constellation leads to back up opposition of

= ( 12 * 3 ) T / ( 4.5 * 1.5 ) M2

= 5 t/m2

Figure 3: support design for galleries

Rock burden at junctions ( BY CMRI-RMR )

Rock burden at junction of gallery and split in depillaring countries has been estimated utilizing the undermentioned empirical equation of CMRI-RMR system:

Rock burden = 5 A- B0.3A- D ( 1-RMR/100 ) 2aˆ¦aˆ¦aˆ¦.. ( 3.3 )

Rock burden = 3.17 t/m2

Design OF SUPPORT SYSTEM

To back up the stone burden we need to give support higher than the stone burden at that country so that the roof will non fall

Its assumed that junctions will be supported by cement grouted bolt ( 6 ton ) & A ; wedges

Junction dimension = 4.5 * 4.5 m

Roof bolt capacity = 6 T

Bolt spacing = 1.5 m

Distance between two rows of bolts = 1.5 m

Thunderbolts in in-between row = 3 & A ; in another two row of support wedges with 1 bolt in between.

So entire bolt =5 & A ; wedge = 4 as given in the figure4

The above constellation leads to back up opposition of = { ( 4*30 ) + ( 5*6 ) } / ( 4.5*4.5 )

=150 t/20.25m2

= 7.4 t/m2

Safety factor = 7.4/3.17

= 2.33

Figure 4: support design for junctions

ANOTHER Approach

Its assumed that junctions will be supported by full column rosin bolt

Roof bolt capacity = 12 T

Bolt spacing = 1.5 m

Distance between two rows of bolts = 1.5 m

No of bolts in a row= 3

No of row = 3

The design shown in figure 5

The above constellation leads to back up opposition of 108 t/20.25sq.m = 5.33 t/sq.m

Safety factor = 5.33/3.17

= 1.68

Figure 5: design of support system for junctions

3.4 ESTIMATION OF ROCK LOAD & A ; DESIGN OF SUPPORT SYSTEM IN ( depillaring workings )

3.41 Rock Load in Slice and Goaf Edge by ( NGI-Q SYSTEM )

Rock burden ( Proof ) in piece and goaf border was estimated utilizing NGI-Q system from the undermentioned empirical relation:

Proof = 2/3 ( Jn1/2/ Jr ) x ( 5Q ) -1/3aˆ¦aˆ¦aˆ¦aˆ¦aˆ¦ . ( 3.4 )

Where, Jn = 9, Jr = 1.5, Q = 2 for piece and Q = 1 for goaf border.

Hence, stone burden in piece, Proof is 6.19 t/m2 ; and sway burden at goaf border, Proof is 7.79 t/m2.

Design OF SUPPORT SYSTEM FOR ( pieces and goaf borders )

It ‘s assumed that piece & A ; goaf borders will be supported by steel props & A ; wedges.

Slice width = 5m

Rock burden in piece, Proof is 6.19 t/m2

Steel prop capacity = 30 T

Prop spacing = 1.5 m

Distance between two rows of prop = 1.2 m

Chock capacity = 30 ton

In 2.4m of length support system will be 1 wedge & A ; 3 steel prop as shown in figure. Goaf edge side will hold 1 wedge & A ; 1 prop with spacing 1.2 m. the other side will hold 2 props with spacing 1.2m.

Figure 6: Design of support system for pieces & A ; goaf borders

The above constellation leads to back up opposition of 120 t/12sq.m = 10 t/sq.m

Safety factor = 10/6.19

= 1.61 for pieces

For goaf borders

1 wedge & A ; 1 prop are at that place in 2.5m breadth & A ; 2.4m length so

The support opposition is = ( 30 + 30 ) / ( 2.5*2.4 )

= 60 / 6 = 10t/m2

Safety factor = 10 / stone burden for goaf border

= 10 /7.79

= 1.28

It is good because goaf border is supported for impermanent period.

3.5 NUMERICAL MODELLING STUDY

The 3 pillars have been modelled utilizing FLAC5.0 with 4 galleries.

The mold process is

1. OPEN FLAC 5.0

2. Travel to GRID & A ; do grid

3. so travel to ASSIGN to do galleries & A ; give stone belongingss

Travel to STRUCTURE to back up gallery with stone bolt or overseas telegram bolt or props.

Then travel to run for computation of FOS

Travel to PLOT to see different profiles

Figure 7: grids demoing four galleries

Figure 8: bolting in galleries ( tap colour 1s are bolts )

After giving proper support solve it and see the factor of safety.. & A ; emphasis & A ; strain distribution..

Chapter 4

Analysis AND RESULTS

Figure 9: emphasis distribution around galleries

Figure 10: strain distribution around galleries

From the figures we can see that the emphasis distribution is more concentrated in the center of the pillar.

There is no formation of tenseness zone, so the 3 overseas telegram bolts are sufficient to back up a gallery.

Chapter 5

CONCLUSION AND RECOMMENDATIONS

For gallery

To back up gallery, Resin bolt capacity = 12 T, Bolt spacing = 1.5 m, Distance between two rows of bolts = 1.5 m, No of bolts in a row= 3 is required which gives a safety factor of 2 which shows the support is reasonably plenty for the gallery.

For junctions

To back up junctions, Resin bolt capacity = 12 T, Bolt spacing = 1.5 m, Distance between two rows of bolts = 1.5 m, No of bolts in a row= 3, No of row = 3 is required to back up junction which gives a safety factor 1.68. It shows the junction is good supported & amp ; there is no fright of roof autumn or Roof bolt capacity = 6 T, Bolt spacing = 2m Distance between two rows of bolts = 2m Bolts in in-between row = 3, and in other 2 row 1 each in between wedges so entire = 5 roof bolts Chocks = 4 at each corner Can back up the junction with safety facto 2.33. It shows the junction to a great extent supported & A ; applied when the junction roof status is really bad.

Slices & A ; goaf borders

In 2.4m of length support system will be 1 wedge & A ; 3 steel prop. Goaf edge side will hold 1 wedge & A ; 1 prop with spacing 1.2 m. the other side will hold 2 props with spacing 1.2m.

The support system gives a factor of safety 1.61 for pieces & A ; for goaf border side it gives 1.23 of safety factor. Goaf edge side safety factor less it is good for undermining & A ; we even do n’t necessitate more support at that place, as no individual is employed at that place to work.

Flac5.0

By the mold we saw that for those given status, 3 overseas telegram bolts are more than sufficient to back up the gallery.