opencast stability

7/28/2019 Opencast Stability

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23Vol 87 , August 2006

Dr I Roy is with the Central Mine Planning and Design Institute

(CMPDI), Ranchi 834 001.

This paper (modified) was received on January 10, 2006. Written discussion onthe paper will be received until October 30, 2006.

Effect of Opencast Mine Floor Inclination on Stability of

Back-filled Dump under Varying Hydro-geological Conditions

a Case Study

I Roy,Member

Dump/mine floor inclination is an area in which there is no control of mine planners and operators though inclination of dump floordictates the stability of back-filled dump to a larger extent. Due to continuous formation of back-filled dump and tremendous pressure ofcoal production on the mine operators, levelling of dump floor is not possible unlike civil engineering embankment. The subject of specialinterest in this article is to discuss the methodology of stability calculations with emphasis on the influence of mine floor inclination in thestability of back-filled dump. The effect of varying height of ground water table above the dump floor is also discussed in this article. Thepresentation of the article is in two parts. In the first part, the methodology of stability analysis is discussed in brief. The second part dealswith a specific case study of a mine, namely, Dhanpuri Opencast Coal Mine, India which is having abrupt variation in mine floor inclinationalong with variation of height of ground water table within the mine property.

Keywords: Dump slope stability; Geo-technical engineering; Opencast mining; Reclamation; Waste management

INTRODUCTION

Mine planners and operators are always in a difficulty to decide the

maximum allowable height and safe slope angle of back-filled

dump in an inclined mine floor. Allowable dump height and

corresponding slope angle for a stipulated factor of safety varies

from one sector to another of the mine for a change in mine floor

inclination or hydro-geological conditions when there is no change in

other geo-engineering parameters. The geo-engineering parametres1

considered in this investigation are as follows.

Geo-technical parameters of

(i) dump material, and

(ii) interface material between dump and foundation (Figure 1)

(after mining of coal, a layer of crushed rock and coal dust

submerged under water lies on the floor of the mine in slushy

condition which is termed in this study as interface material).

Dump/mine floor inclinations

Hydro-geological parameters like

(i) upward thrust of water due to accumulated water table

within the dump, and(ii) seepage force of water due to accumulated water table

within the dump.

Effect of seismic force on the dump mass.

Coal rib / barrier dimension and shear strength parameters of

interface material between coal barrier and its floor (Figure 1).

This article presents methodology of stability analysis of back-filled

dump. This article also presents a case study2 showing the influence

of the followings :

(i) dump/mine floor inclination, and

(ii) ground water table variation on stability of back-filled dump

of Dhanpuri Opencast Mine (located in central part of India).

METHODOLOGY OF STABILITY ANALYSIS

A computer program has been developed using the following

computational methods for stability analysis of back-filled dumps1.

Back-filled dump in opencast coal mining operation experiences twotypes of failure surfaces3 (Figure 1 and Figure 2), namely :

(i) circular failure surface, and

(ii) circular-cum-planar failure surface4.

Following sequence of computations is envisaged in stability analysis

of back-filled dumps [Figure 1].

Determination of factor of safety of first trial surface ABDC

(circular-cum-planar failure surface) by Fellinius Method [ Figure 2].

An iteration method to locate most critical failure surface

corresponding to absolute minimum factor of safety by

Fellinius Method1 [Figure 3].

Absolute minimum factor of safety corresponding to most

critical failure surface by Fellinius Method is further modified by

Bishop's Simplified Method.

The factor of safety by Bishop's Simplified Method is adjusted

comparing with stipulated value of factor of safety to achieve

safe and economic combination of back-filled dump profile.

Factor of Safety by Fellinius Method [Figure 2]

The trial failure surface ABDC is divided into suitable number of

sectors, such as, BDFG and it is further divided into fifty number

of slices (J=1,50) of equal thickness. The methodology of stabilitycalculation for one of the slice PQRS is documented below (Figure 2).


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24 IE(I) JournalMN

Each slice is subjected to its own weight and other forces like asfollows.

Upward thrust of water due to accumulated water table withinthe dump5.

Seepage force of water due to accumulated water table withinthe dump .

Seismic force on the dump mass.

All these forces acting on each slice [PQRS] are resolved into thecomponents to determine disturbing and frictional force onindividual slice. They are suitably added for all slices respectively todetermine the cumulative disturbing and frictional forces acting onthe entire trial surface. The above forces for slice PQRS are determinedas documented below

W(J) = [H1(J)+H1 (J-1)] 0.5TkDWT. (1)

where W(J) is the weight of slice PQRS; Tk, the width of slice; D,the bulk density of dump mass under natural moisture conditions;

WT, the width of failure block in strike direction, and 0.5{H1(J)+H1 (J1)} is the average height of the slice.

Upward thrust and seepage force of accumulated water on thedump mass is considered in the following manner.

Phreatic surface of water {DPX(J),DPY(J)} is determined byfollowing Casagrande's equation.

DPY(J)=[(DPX(J) PacosL/2)2PasinLtanL]0.5 (2)

where Pa is the length of oozing of water along the slope of dump,that is, inclined plane AF [Figure 2]; Hw, the height of accumulated

water table with respect to horizontal plane passing through toe ofdump and Lis the slope angle of dump.

Upward thrust of water can be calculated as

{DPY(J)(N(J)+N(J-1))/2.0}TkDwWT (3)

Seepage force [V(J)] can be calculated as

Tk{DPY(J)(N(J)+N(J1))2.0}WTDwtan[X(J)] (4)

Circular cum planner surface

Circular surface

E(3, Y3)

Slope angle, L

Dip direction

A (0,0)

Pa

Z

GF

B

Sector BDFGInterface material

PQRS

D

C

Accumulated water table

HJ

Width of failure mass (WT) instrike direction

HwH

P

Q

U

Seismic force

Disturbingforce

S

Upward thrust resisting force

Tk

R Y(J)

X(J)Seepa

ge force

Slice PQRS

W1(J)

Co-ordinates of R = M (J), N(J)Co-ordinates of S = M (J-1), N(J-1)

H1(J)

DPPX (J), DPPY (J)

DPX (J), DPY (J)

Figure 2 Factor of safety calculation by Fellinius and Bishop's simplified method

230

220

210

200

190

180

170

160

150

140

130

120

110

100-130 -120 -110 -100 -90 -80 -70 -60 -50 -40 -30 -20 -10 0.00 10 20 30 40 50 60 70 80 90 100 110 120 130 140 150 160 170 180 190 200 210 220

Waste rock dump by haul truck

Waste rock dump by dragline

Interface material

4m

10m

Berm at coal rib level To be rehandled (abcd)

Berm at dragline placement level

d

Shoval and truck working level

Dragling working level

ba

Circular failure

Circular cum planar failure

Coal rib/barrier

Soil strata

Rock strata

Coal

Mine advancing towards dip direction

Figure 1 Schematic representation of dump formation in a drag-line opencast mine

(Not to scale)


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where Dw is the density of water andX(J) is the gradient of phreatic

surface/ ground water table

tan{X(J)}=[{DPPY(J)DPPY(J1)}{DPPX(J)DPPX(J1)}] (5)

whereX(J) is the gradient of accumulated water table or phreaticsurface.

Co-ordinates of T: DPPX(J), DPPY(J)

Co-ordinates ofU: DPPX(J-1), DPPY(J-1)

Considering upward thrust of water, the weight of slice W1(J) can

be calculated as

W1(J) = W(J) Upward thrust of water (6)

where co-ordinates of base point of PQRS, that is,Ris denoted by

M(J), N(J).

Considering all the above forces, disturbing and resisting force is

calculated in following manner.

Disturbing force

= W1(J)sinY(J)+AW1(J)cosY(J)+V(J)cos{Y(J)X(J)} (7)

where A is the seismic co-efficient of surrounding area of the

Dhanpuri Mines, and Y(J) is the gradient of base plane RS.

Frictional force = [W1(J) cos Y[J]A W1(J) sinY(J)V(J)

sin{Y(J)X(J)}] tan(angle of internal

friction of dump/interface material).

Cohesive force in slice PQRS is determined in following manner.

Surface area of failure mode = (RSWT)

Curved length of RS = (Radius of the failure surface)(Z in radian

within slice PQRS).

Cohesive force = Surface area of failure mode(Cohesion of dump

mass). (9)

Cohesive force within sector ABG is determined in following

manner.

Cohesive force = Surface area of AB Cohesion of interface

material. (10)

After mining of coal, a layer of crushed rock and coal dust submerged

under water lies in Indian coal mines which is termed in this study asinterface material. Hence, cohesion and angle of internal friction of

interface material is determined in the laboratory after keeping the

material in submerged condition.

Resisting force = Frictional force + Cohesive force +Resisting force

due to coal rib (11)

The ratio of the cumulative resisting force of all the slices within

failure surface to the cumulative disturbing force is the factor of safety

of the selected trial surface ABDC by Fellinius Method.

Iteration Method to Locate Most Critical Failure Surface

[Figure 3]

The above method of factor of safety determination, as discussed

earlier, is repeated by selecting in a systematic manner several trial

surfaces around the first one to locate the most critical failure surface

through a generalized iteration scheme [Figure 3] developed in this

investigation as documented below.

The iteration scheme is initiated by selecting any point 0, as the centre

of first trial failure surface. Therefore, the iteration scheme is adoptedto search around this point and locate the point, at which the factor

of safety is the absolute minimum. To achieve this situation,

following steps are followed in the sequence as given below.

Step 1

Having selected the point 0 as the centre of the first trial failure surface

passing through toe of the dump (A), the value of the factor of

safety(FS1) of the first trial failure surface passing through toe of the

dump(A) is determined.

Step 2

The point is next changed in the vertical direction RR to the nextpoint 0x and the value of the factor of safety(FS

2), for the

corresponding failure surface is determined.

Step 3

If the value of the factor of safety(FS2) with centre 0x, is found to be

less than the previous value of the factor of safety(FS1), next point is

selected, beyond this point, along the same direction(RR). The

procedure is repeated, at regular intervals along the same vertical

line(RR), till the point, at which the value of the factor of safety

increases from the previous one (FSm) is located. This above-

mentioned previous value of factor of safety(FSm) is used for

subsequent checking along the next line(RR1). This procedure isrepeated along all the other lines, at regular intervals of delta. The

point 01, corresponding to which, the value of the factor of safety, is

the real minimum, is thus located in this zone.

Step 4

Having thus located the centre of the toe circle 01 for the next

iteration, this point 01 is adopted as the starting point and steps 1

through 3, as described above, are followed to search and locate the

point 02, around the zone-01, which gives the minimum of all the

factor of safety probed so far, including the previous zone around 0.

Step 5

If the linear differences between the location of the centres thus located

in the two zones around 0 and 01 represented by 0-01 and 01-02,

respectively, are not within the acceptable accuracy, the iteration

scheme continued to subsequent zone 03, as described in step 4, till

the desired accuracy of the numerical evaluation is achieved.

Step 6

The above process is repeated for trial circles intersecting floor away

from the toe of the dump(X1=0,-10,-20,-30 and so on) and the

factor of safety of failure surfaces passing through toe and away

from toe are compared to find the absolute minimum factor of

safety and corresponding failure surface [0n is the centre of failuresurface with absolute minimum factor of safety (FSam)].

(8)


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26 IE(I) JournalMN

Factor of Safety by Bishop's Simplified Method

As the factor of safety (FS1= FSam) determined by Fellinius Method

gives some error, though on lower and safer side, Bishop's

Simplified Method is applied to determine fairly accurate results in

the following manner [Figure 2]

FS2=[Cohesive force + W1(J) tan (angle of internal friction of

dump/interface material)] M[W1(J) sin{Y(J)} +A W1(J)

cos{Y(J)} + V(J) cos{Y(J)X(J)}] (12)

M= [1+{tan{Y(J)} tan(angle of internal friction)} FS1] cos {Y(J)}

(13)

FS3 = [Cohesive force + W1(J) tan (angle of internal friction)]

M[W1(J) sin{Y(J)} +A W

1(J) cos{Y(J)} + V(J) cos{Y(J)X(J)}]

(14)

M= [1 + {tan{Y(J)} tan (angle of internal friction)} FS2]

cos {Y(J)}

Similarly, by repetitive convergence method FSnwill be equal to FS

n-1.

where FSn is the factor of safety by Bishop's Simplified Method.

Determination of Safe and Economic Combination of

Height and Slope of Dump Profile

The factor of safety by Bishop's Simplified Method (FSn) is

compared with the stipulated value of factor of safety. If the

calculated factor of safety is less than the stipulated value, then the

slope is to be flattened. Otherwise, if the calculated factor of safety is

more than the stipulated value , then the slope is to be steepened sothat factor of safety is equal to 1.2.

Factor of Safety

After considering all the recommended factor of safety6 suggested bydifferent agencies, such as, 'National Coal Board ,U.K, United States

D' Appolonia Consulting Engineers, Mines Branch Canada, Stabilityof Pit Slopes and Dumps by G L Fiesenko, Russia for surface mineslope design, a factor of safety of 1.20 is envisaged in this study.

CASE STUDY

Dhanpuri Opencast Mine of South Eastern Coalfield , located incentral part of India is divided into two parts, that is, Eastern and

Western parts(Figure 4). The mine is characterized by an abruptchange in dump/mine floor inclination from 3o to 8o . There is anaccumulated water of 20 m in the western part of the mine and

very negligible in the eastern sector.

The geo-mining parameters of the mine are as follows2.

Maximum height of the back-filled dump', m : 60 - 70.

Dump floor inclination, Degree : 3 - 8.

As the mine is situated in the seismic zone-I of India, horizontalseismic co-efficient of 0.01 is considered in this study for stabilityanalysis. The floor of the mine is covered with a slushy mixture ofcrushed rock and coal dust submerged under water. This material istermed here as interface material. Geo-technical properties of dumpand interface material is determined in a reputed laboratory of Indiain the following manner.

01, 02, 03......0n-1 : Centres of circles having the least factor ofsafety of the trial circles selected around 0,01,02...0n-2,respectively.

On-centre of most potential failure surface with absoluteminimum factor of safety

ABC : Failure surface which comprises of AB and BC.AB : Planar failure surface through interface materialBC : Circular failure surface through dump material

0 x - Second centre of failure surfaces

RR

RR1

0n

02

01

Zone for centres ofcritical circles

270o

90o

0X

0

Delta = 36 degree (360/20)as there are 10 lines aroundpoint 0.

Delta

H

Dump mass

C

BA

X1

X1 = 0, -10, -20 Interface material at dump floorFigure 3 Iteration method to locate most critical failure surface


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Laboratory Testing

The dump material of around 6-cm size was pre-compacted at anaverage stress of whole dump mass, that is, for a height of 60-mdump mass with an average bulk density of 20 kN/m3, the averagepre-compaction pressure was 600 kN/m2 [30m20 kN/m3]. Then

the dump material was subjected to direct shear test in a large shearbox [30cm 30cm] for different normal stresses prevailing atdifferent levels of dump mass. The dump materials and interfacematerials were tested under natural moisture condition andsubmerged condition, respectively for representation of actual siteconditions. The value of cohesion determined in the laboratory was

verified with the value of cohesion determined by back stabilityanalysis of an existing dump standing at limiting equilibrium ( thatis, total disturbing force = total resisting force) using the laboratorydetermined value of angle of internal friction. The strengthparameters adopted for stability calculations are given in Table 1.

Management of opencast coal mines in India applies a new technique

of leaving a very small part of coal seam/strata at the toe of thedump(which is termed here as coal rib/ barrier) [Figure 1]. In thisproject , the present practice is to keep a coal rib with minimum widthof 10-m at the base and 4-m at the top.

Cohesive resistance and frictional resistance per 10-m width in strikedirection (that is, WT=10-m) [Figure 2] at the bedding planebetween coal rib and its floor is calculated as 20 000 kN and 10 227 kN,respectively taking into consideration following parameters of coalrib [Figure 2].

Base width of coal rib, m : 10.

Top width of coal rib, m : 4.

Height of coal rib, m : 14 ( full coal seam thickness)

Width(WT) of dump failure mass in strike direction [Figure 1], m : 10.

Bulk density of coal, kN/m3 : 24.

Cohesion at the bedding plane between coal rib and its floor, kN/m2 : 200.

Frictional angle at the bedding plane, Degree : 23.5.

The geo-technical parameters of bedding plane which is fractured

due to blasting in coal are determined from drilled core throughbedding plane.

RECOMMENDATION

Safe slopes for different heights of back-filled dumps consideringall the above parameters with factor of safety of 1.2 are calculated

and documented in Table 2 - Table 5.

To explain the application of computer simulated programme, inputparameters considered for stability calculation of 70m high dump(Table 2) is documented below

Height of dump (H), m : 70

Slope angle of dump with respect to dump floor (L), Degree : 26

As discussed earlier, the slope angle of 26o

corresponding to 70-mhigh dump is obtained by trial and error method in such a mannerthat the factor of safety becomes equal to 1.2.

Dump floor inclination (I), Degree : 3

Geo-technical parameters considered as mentioned in Table 1 aregiven below

Pa (length of oozing of water along the slope of dump) as observedin case of Western sector with dump floor inclination of 3

o= 7-m

[Figure 2 and equation (2)].

Seismic co-efficient (A) as per seismic zone map of India = 0.01[Figure 2 and equation (7)].

Cohesive and frictional resistance (that is, WT=10-m) [Figure 2] atthe bedding plane between coal rib and its floor is calculated as 20 000 kNand 10 227 kN, respectively.

Method of Mining for Maintaining the Recommended Slopes

In Dhanpuri Opencast Coal Mine, rock overburden above coal

Table 1 Geo-technical parameters of materials

Strength Dump Materials in Interface Material in

Parameters Natural Moisture Submerged Condition

Condition

Cohesion, kN/m2 80 30

Angle of internalfriction, Degree 22 13

Bulk density, kN/ m3 18 20

Table 2 Details of safe slope of the Western Sector

( Ground water table, m : 20, Dump floor inclination, Degree : 3)

Height of Dump, m Slope of Dump with Respect toHorizontal, Degree

70 29

60 32

Table 3 Details of safe slope of the Western Sector

( Ground water table, m : 20, Dump floor inclination, Degree : 8)


70 26

60 28

Table 4 Details of safe slope of the Eastern Sector

( Ground water table, m : 0, Dump floor inclination, Degree : 3 )


70 31

60 34

Table 5 Details of safe slope of the Eastern Sector

( Ground water table, m : 0, Dump floor inclination, Degree : 8 )


70 27

60 30


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External dump

Final surface boundary of the Mine/Quarry

Present surface boundary of the Mine/Quarry

+ 475mSite for collection of

dump material samplesfor laboratory testing

Dip

ofth

eMin

e

Strike

direction

+ 480m

Back-filled dump(Western Sector)

+ 485m

450m

420m

Back-filled dump(Eastern Sector)

Present floorboundary of the

Mine/Quarry

Figure 4 Working plan of Dhanpuri Opencost Mine

seam/strata are blasted by explosives and excavated by followingtwo techniques [Figure 4].

Overburden Rock Removal by Dragline

The dragline excavates the blasted rock and dumps the fragmentedrock immediately in the earlier de-coaled area. The dragline also re-handles some portion of waste rock (ABCD)[Figure 1] which needsto be kept minimum considering the economics of the mine.

Overburden Rock Removal by Shovel Truck Combination

The shovel excavates the rock and dumps back to the haul truck. Thehaul-truck moves along the haul road and dumps the wastefragmented rock above the dragline dump [Figure 1].

The above recommended slopes of dump are maintained byincreasing/decreasing following three parameters of dump profile[Figure 1].

Berm width at the dragline placement level.

Berm width at the coal rib level.

Slope angle below dragline sitting level.

The re-handling volume (ABCD) dictates the above threeparameters. But the amount of re-handling also needs to be keptminimum considering economics of the mine. Hence, acompromise is to be made by mine management in optimising the

volume of re-handling (ABCD) so that overall slope is within thepermissible limits as mentioned in Table 2 - Table 5.

Other than maintaining the above-mentioned recommended slopes,the interface material (that is, the water submerged layer of wasterock and coal dust) is recommended to be cleaned as far as possible tomake the mine/dump floor competent.

CONCLUSION

Following inferences are drawn based on the above study.

Dump slope stability analysis should be made mandatoryfor all the major opencast mines in India.

In Dhanpuri Opencast Coal Mine, there is an abruptvariation in inclination of mine floor and also variation ofaccumulated water table within the dump mass. As evident

from Table 2 - Table 5, both of the above parameters play animportant role in the stability of dump. Hence, in everyopencast coal mine of India, separate dump profile (heightand slope) should be followed for different mine floorinclinations and for different position of accumulated watertables within the back-filled dump mass.

Optimization of height and slope of dump for a stipulatedfactor of safety should be carried out by stability analysisconsidering all the geo-engineering parameters like geo-technical, hydro-geological and seismic parameters .

The safe and economic combination of dump profile can bemaintained by optimizing the re-handling volume (ABCD)as shown in Figure 1. To reduce the overall slope angle of


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dump profile, the re-handling volume by dragline is to beincreased. But, the re-handling volume is to be keptminimum as far as possible considering the mineeconomics. Hence, the stability analysis helps in deciding theoptimum volume of re-handling by dragline from bothsafety and economical point of view.

ACKNOWLEDGMENT

The author is thankful to South Eastern Coalfields Limited, Indiafor sponsoring this study to Central Mine Planning and DesignInstitute, India and for providing necessary facilities and informationduring the course of this study.

REFERENCES

1. I Roy. 'Influence of Geo-engineering Parameters on the Stability of

Dumps, (PhD Dissertation)'. Indian Institute of Technology, Kharagpur, India,

1999.

2. 'Consultancy Report on Internal Dump Stability of Dhanpuri Opencast

Mine, South Eastern Coalfield Limited, India'. Central Mine Planning and Design

Institute, 2003.

3. K E Hebil. 'Spoil Pile Stabilisation at the Paintearth Mine, Forestburg,

Alberta'. 'International Symposium on Geo-technical Stability in Surface Mining,

Calgary, Canada, 1986, p 181.

4. D B Cambell. 'Stability and Performance of Waste Dumps on Steeply

Sloping Terrain'. International Symposium on Geo-technical Stability in Surface

Mining, Calgary, Canada, 1986, p 317.

5. G W Wilson. 'Embankment Hydrology and Unsaturated Flow in Waste

Rock'. Society for Mining, Metallurgy and Exploration, 2001, p 305.

6. W L Gerhard. 'Steepened Spoil Slopes at Bridger Coal Company'. Society for

Mining , Metallur gy and Exploration, 2001, p 351.

7. S K Das. 'Problems of Highwall and Spoil Dump Stability and Various

Preventive Measures in Highly Mechanised Opencast Mines'. The Indian

Mining and Engineering Journal, 2001, p 63.

opencast stability

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