UNIVERSITY OF MORATUWA

Faculty of Engineering

DESIGN REPORT

MODULE ER4122: DESIGN PROJECT

DASSANAYAKE D.M.S.M. (100855B)

PUSHPAKUMARA K.B.N. (100850F)

SAMPATH R.P.S. (100836T)

PRABHANGA U.B. (100830U)

Department of Earth Resources Engineering

Date of Submission 02-12-2014

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PREFACE

The major goal of designing a mine utilizing the optimum strategies is the gaining of economic

stability when the hour of need arrives. No mine in the known world can be designed for 100%

safety and overwhelming profits. Where there is a mine there is a threat for the people who are

working in there, there is a threat of major economic draw backs and there is a threat for the

natural habitat and the environment.

The ultimate objective of a mining engineer or the group of design engineers is the minimization

of threats and weaknesses and maximization of the strengths and opportunities, the desired mine

will be facing in the future, past and most importantly in the present. This is the purpose of this

whole design project. This particular design of the mine is the ultimate design we came up as a

group for our final year design project.

In the first chapter, an introduction is given regarding the location of the mine and the available

details for the designing. In this context, the method which we used to determine the dip and strike

of the ore body and the three dimensional visualization of the area geology are presented briefly.

In the second chapter exploration program is discussed. Under that, the exploration done using

the diamond drilling program and the selection of the locations for drilling are discussed. We

strictly followed the continuity of the report, at the same time answering the desired questions

which were attached with the initial design project materials.

In the third chapter, the mining method and the operational and strategic utilization of the

engineering knowledge in designing the mine are expressed. The design regarding the

underground mineral transport system is also elaborated with general details.

Final chapter is allocated for the costing and economical evaluation of the mine. With the

available information we could build up a systematic approach to justify the accountability of the

whole design project.

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Table of Contents

PREFACE .......................................................................................................................................................... 2

1. CHAPTER I ................................................................................................................................................ 6

1.1 Introduction ..................................................................................................................................... 6

1.2 Calculating the strike and dip of the veins ...................................................................................... 7

1.3 Surface distribution of the veins ............................................................................................................ 8

1.3 Exploration ...................................................................................................................................... 9

2.1.1 Diamond Drilling Programme ........................................................................................................ 9

2.1.2 Developing access roads............................................................................................................... 14

1.4 Mine opening ................................................................................................................................. 16

2. CHAPTER III ............................................................................................................................................ 18

3.1 Addit .................................................................................................................................................... 18

3.1.2 Drilling pattern of the Addit ............................................................................................................. 18

3.2 Blind Shaft ........................................................................................................................................... 20

3.2.3 The auxiliary shaft of the mine ..................................................................................................... 21

3.2.4 Blind shaft & auxiliary shaft sinking ............................................................................................. 22

3.3 Shaft lining ........................................................................................................................................... 24

3.5 Mining method .................................................................................................................................... 28

3.5.1 Overhand cut and fill method ....................................................................................................... 28

3.6 Mine Layout ........................................................................................................................................ 31

3.7 Underground transportation system .................................................................................................... 32

3.8 The conveyor belt system .................................................................................................................... 34

3.8 Ventilation methods ............................................................................................................................ 36

3.8.1 Stope ventilation methods ............................................................................................................ 36

3.9 All the costing has attached to appendixes

3. REFERENCES .......................................................................................................................................... 38

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List of figures

Figure 1.1 The contour map of the area .................................................................................................................. 6

Figure 1.2 The projected strike lines of the veins ................................................................................................... 7

Figure 1.3 The graphical representation of the dip and strike ................................................................................. 7

Figure 1.4 The expected surface distribution of the veins ....................................................................................... 8

Figure 1.5 The 3D visualization of the vein ............................................................................................................ 8

Figure 2.1 The hypothetical pattern of the exposed point distribution .................................................................. 10

Figure 2.2 The horizontal depth to 500 m depth of the ore ................................................................................... 10

Figure 2.3 The 3D visualization of Drilling lines from the drilling point ............................................................. 11

Figure 2.4 The drilling locations ........................................................................................................................... 12

Figure 2.5 The map of the access roads ................................................................................................................ 14

Figure 2.6 Adit opening......................................................................................................................................... 16

Figure 3.1 The proposed addit with its dimensions ............................................................................................... 18

Figure 3.2 Drilling pattern and the blasting sequence of the addit ........................................................................ 19

Figure 3.3 The location of the blind shaft ............................................................................................................. 20

Figure 3.4 Blind Shaft features and dimensions .................................................................................................... 21

Figure 3.5 Auxiliary shaft Dimensions ................................................................................................................. 22

Figure 3.6 Main shaft & Auxiliary shaft blasting pattern ..................................................................................... 23

Figure 3.7 The winch and the steel structure ......................................................................................................... 24

Figure 3.8 The cross cut with dimensions ............................................................................................................. 25

Figure 3.9 The blasting pattern of the cross cut .................................................................................................... 27

Figure 3.10 The ore blocks created to be mined ................................................................................................... 28

Figure 3.11 The layout of the mine with cross cuts, winzes and shafts ................................................................ 29

Figure 3.12 The extraction procedure ................................................................................................................... 30

Figure 3.13 3D view of the mine lay out ............................................................................................................... 31

Figure 3.14 The dimensions of the rail cross section ............................................................................................ 32

Figure 3.15 The mine car with dimensions ........................................................................................................... 33

Figure 3.16 The deflection of the pulley ............................................................................................................... 34

Figure 3.17 Pulley width ....................................................................................... Error! Bookmark not defined.

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List of tables

Table 1-1 The Dip and Strike of the veins .............................................................................................................. 8

Table 2-1 Distances to the drilling position .......................................................................................................... 11

Table 2-2 The drilling length from the drilling locations ...................................................................................... 13

Table 3-1 The location of the blind shaft .............................................................................................................. 20

Table 3-2 Factor of safety for the rails .................................................................................................................. 33

Table 3-3 Pulley width recommendations ............................................................................................................. 35

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1. CHAPTER I

1.1 Introduction

This report is comprised of the mine design done by the group members of the final year design

project group. The mine described here is a hypothetical mine. According to the given information

the area geology suggests a steeply dipping graphite mineralization. The contour map including

the mineralization of the area was given. The contour map is shown in figure 1.1

Figure 1.1 The contour map of the area

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1.2 Calculating the strike and dip of the veins

The contour interval was given as 70m. Therefore we can generate the lines of strike of different

levels as shown in figure 1.4.

59o

Figure 1.2 The projected strike lines of the veins

Figure 1.3 The graphical representation of the dip and strike

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1.3 Surface distribution of the veins

Surface distribution pattern of the veins considering the vein thickness as 40cm and the marked

points are in the center of the vein.

Figure 1.4 The expected surface distribution of the veins

Figure 1.5 The 3D visualization of the vein

Vein Dip Strike

C 59o

9o 29 24

D 63o

181o 51

Table 1-1 The Dip and Strike of the veins

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CHAPTER II

1.3 Exploration

2.1.1 Diamond Drilling Programme

The need of the diamond drilling arises when the sub surface distribution of the ore body has to be

identified. It deviates from the earlier exploration methods as it allows the removal of solid

cylinders of rock or often what referred to as core from deep within the earth. The materials

present in the desired depth levels are visually inspected and later detailed investigations are run

on them to identify what actually exists in the given depths.

There are basically two ways in which the diamond drilling programme can be conducted.

1. Inclined core drilling

The main advantage in this method is that the drilling machine can be kept at one

location. The drilling can be done at various directions. Since the angle of drilling

increases the length between the drilling position and the desired ore body identified,

the drilling length is usually high. It, the inclined core drilling method, has to be

selected with great caution. When the estimated mobility cost of the drill machine is

higher than the estimated drilling cost, application of this method is cheaper.

2. Vertical core drilling

The drilling can be done in various locations but the only condition is that the drilling

should be done vertically penetrating the subsurface strata. The drilling length is always

the vertical length and no any alteration regarding the length can be achieved other than

changing the drilling location. When the estimated mobility cost of the drill machine is

higher than the estimated drilling cost, application of this method is costlier.

In this particular design-exploration programme, the depth at which the exploration ends is

defined as 500m. Since the area geology is asked to be assumed closer to the one that in Bogala

area (i.e. rock type; Metamorphic garnet biotite gneiss), with the dip of the mountain (higher cost

for mobilization and stabilization of the drilling machine), the best way to do the diamond drilling

is inclined core drilling.

The requirement of the diamond drilling program in this context is to explore the underlying

geology. Therefore out of the two methods of rotary drilling and wireline drilling, wireline

drilling method is chosen. In wireline drilling method, it allows the inner tube of the core barrel

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to be raised to the surface and empty the barrel to be pumped back into place without removing

the drill string or diamond tools from the hole. Because of that there are long distances to be

drilled this method is more suited.

Figure 1.6 The hypothetical pattern of the exposed point distribution

Let the distance between two parallel lines going parallel to the strike is X, the number of times

which X has to be multiplied in order to find a location on the surface which meets the vein at

500 m depth on the line going perpendicular to the strike is c. The dip angle is .

c x X

500m

c x X tan = 500 m

Figure 1.7 The horizontal depth to 500 m

depth of the ore

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The distances are calculated to the point where the ore body will be lying 500 m below.

X c Horizontal Distance

to the Location

Vein C 36 m Vein C 59o

Vein C 8 Vein C 288 m

Vein D 41.99 m Vein D 63o

Vein D 6 Vein D 249 m

Table 1-2 Distances to the drilling position

But based on the above calculations the drilling points will be located way beyond the perimeter

of the given map. Assuming the map is the only reference we can have and there may be other

constructions beyond the perimeter another method has to be developed.

Lets find the point located at equal distances to the exposed vein-c points on the plane of the ore

body.

Inclined ore body

Angled drill holes

Figure 1.8 The 3D visualization of Drilling lines from the

drilling point

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Figure 1.9 The drilling locations

The drilling locations are selected so that the distances from the drilling point to the respective

exposed points are equal. An exception had to be made for the c-vein exposed points due to the

presence of the river. There, drilling point and two exposed points are selected to measure the

distance in each two cases avoiding the buffer zone of the river. That is the reason for the presence

of two C-drill points named, C-drill point1 and C-drill point.

AQ (Hole diameter: 48mm) (refer appendix i) drill bit is selected to be used for the drilling

since, drilling lengths exceeding 1000 m can be achieved by it with minimum failures. Its core

diameter is less and the torque needed to rotate the bit is less and fuel consumption is also less

than other types. The drilling rig that fits the requirements is the hechang-HQY-500 drilling

rig (refer appendix ii).

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Surface drill point location Drilling details

Bearing angle to the

horizontal

axis

Drilling length

D-drill point 1 Directly

downwards

90o 150 m

048o 65

o 200 m

131o

65o

200 m

C-drill point 1 Directly

downwards

90o 150 m

284o 60

0 200 m

206o

80o

200 m

C-drill point 2 Directly

downwards

90o 150 m

325o

65o

150 m

260o

60o

150 m

Table 1-3 The drilling length from the drilling locations

Apart from the initial exploration programme, continuous exploration programmes have to be

conducted while the mine is under operation. The following figure shows the expected locations

and drilling patterns when each level of the mine is under operation.

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2.1.2 Developing access roads

Based on the given area map, two access roads are developed as branches of the main roads

specified in the map. R1 road will be the access road for the vein D addit entrance and R2 will be

the access road for the vein C addit entrance.

In the second stage when the mine is under operation, principle level roads are driven to the

deposit. These roads are rail tracks for the ore transportation purposes. Since the lower levels are

to be accessed, a blind shaft is sunk from the addit and the lower level rail tracks are connected to

this blind shaft. This is the case for both the addits. Hence there are two blind shafts. These two

will be eventually connected to all the cross cuts and auxiliary shaft.

Figure 1.10 The map of the access roads

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2.1.3 Ore grade calculation

When the exploration drilling is over, the shape and the distribution of the ore body can be

visualized. But that doesnt give a clear indication whether the initiation of the mining project is

economically feasible. To find out whether mining is profitable more detailed survey has to be

conducted. The drill core samples can be analysed to find the ore grade. Using statistical methods

the total ore grade of the mine as well as the areas where higher grade deposits present can be

mapped out. If the cut-off grade value is low the parts of the mine where the grade is low can be

mined out quickly, if the cut-off grade is higher some parts of the mine will have to be abandoned.

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1.4 Mine opening

The two veins are located directly opposite to each other. With the river flowing on top of the vein

c it would be difficult as well as costly to use addits to open the ore body. Therefore a better

option would be sinking a shaft at the center of gravity of the two main veins in order to extract

the maximum out of them.

But the circumstance is such that the shaft has to be sunk on the top of the mountain. But that

would require a huge investment. Since the extraction of the ore will not be possible until the

drives are also developed, sinking a main shaft at the beginning would not be a good strategy.

Therefore using one addit to enter into the ore body from the opposite side of the river while an

auxiliary shaft is being sunk from the top of the mountain would be more economically viable

since the revenues will be generated at the initial stage also with the production from the driven

addit. The location of the addit is selected opposite to the river to avoid possible costly operations

such as pumping water from the addit that is being excavated. The first move is merely for

covering up as much as possible for the expenses incurred during the initial mine development

operations.

Figure 1.11 Adit opening

It is expected that the auxiliary shaft will be finished up to the depth of the second level of the

vein D, when the first few levels of the vein D are extracted. Then level by level the vein D can be

extracted using cut and fill method.

When adequate revenues are generated, the extraction can be started from the vein C using the

second addit. The advancing of the addit will open up the first level of the vein C. with that being

developed a cross cut is also developed to meet the auxiliary shaft. So the whole mine lay out

comes in to operation with time.

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When the mine becomes enlarged with the connection of the cross cut from the vein C to the

auxiliary shaft, natural ventilation will not be adequate enough. Therefore an exhaust fan has to be

implanted on the auxiliary shaft.

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2. CHAPTER III

3.1 Addit

The mine is expected to be operated at the beginning using addits. The major reason behind that is

the economic factor. The locations of the ore bodies are such that the safety also will be high if the

initial opening was developed as an addit.

Figure 2.1 The proposed addit with its dimensions

3.1.2 Drilling pattern of the Addit

In this particular blasting pattern, a burn cut at the centre is used to create the space for the burden

of rocks to fall into. The specified blasting sequence helps the fragmented material to pile

systematically without disturbing the flow of the blast. Closer spacing of the drill holes in the

bottom emphasizes the fact that the load is predominant in the bottom part compared to the upper

part of the perimeter.

The following information are assumed in this design.

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Addit area = 4.6 m2

Bit size of the drilling machine =38 mm

Advance of the shaft per round = 1m

Number of holes per round = 53

Powder factor for the blast = 2.7kg /m3

Romanic Numbers half second delay

Arabic numbers m second delay

Figure 2.2 Drilling pattern and the blasting sequence of the addit

Dimension in ft

Cut Hole (64mm)

Drill hole (38mm)

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3.2 Blind Shaft

The entry to the mine is achieved thorough an addit. Therefore to extract the underlying ore body

in the subsurface of the mountain, a blind shaft has to be developed. This shaft is the main

development shaft used to connect the cross cuts which are connected to the ore body from the

other end. Since the shaft is not visible to the outside this is called a blind shaft. To hoist the

materials being extracted from the underground there is a winch located at the top.

Figure 2.3 The location of the blind shaft

The location of the shaft is selected considering two reasons.

1. The closeness of the shaft opening to the river (i.e. the buffer zone of the river)

2. The distances at the subsurface level to the two ore bodies from the shaft (this should be as

minimum as possible fulfilling the condition 1)

Considering all these conditions the most favorable area boundary for the shaft sinking was

identified and the location of the shaft can be given as follows.

location (380, 630)

Elevation (assumed from M.S.L.) 480 to 500 m

Slope 27o

Table 2-1 The location of the blind shaft

Blind Shaft

Addit

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3.2.2 Blind shaft Dimensions & Features

Although this is a blind shaft this may well be acting as a primary shaft in this design as the

mining operations progress. This is the main transport passage into the underground from the

main addit entrance. The 2x7 ft rectangle cross sectional area indicated in the below figure acts as

the passageway for facilitating electricity, compressed air and other necessary supplies which are

continuously needed when the mine is in operation. The skip is used to hoist material as well as

men working underground. To hoist this cage the winch specified earlier is used with wire ropes.

Figure 2.4 Blind Shaft features and dimensions

3.2.3 The auxiliary shaft of the mine

The need of an auxiliary shaft occurs mainly due to two factors.

1. For the emergency exit purposes as a secondary entry path

2. For effective ventilation across the mine.

The dimensioning of the auxiliary shaft is given in the below figure. The shaft is sunk with the

rectangular shape neglecting the resistance created by the shapefactor to the ventilating air flow

considering the fact that the natural ventilation caused by it is high. Since the top most point of the

shaft is located towards the top of the mountain, there exists a natural stream flow development

through the main addit entry towards the upper opening of the auxiliary shaft due the differences

in wind speeds in the respective hights. Thus the natural ventilation is pretty high and the

Skip or Cage

Ser

vic

es f

acil

itat

ing a

rea

Blind shaft

Guide vain

Dimension in ft

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ventillation by fans will be only needed after the first levels of the workings have advanced to a

significant distance .

Figure 2.5 Auxiliary shaft Dimensions

3.2.4 Blind shaft & auxiliary shaft sinking

The whole operation can be emphasized as the following sequence of points.

Preparation for blasting (Cleaning and marking of the blast face)

Drilling

Charging and Blasting

Washing & scaling

Mucking

Supporting

This is a cyclic process. Once the sequence has reached to supporting stage again another blasting has to be

undertaken. It is evident that the presence of the river will cause the water table to rise. When the water

Dimension in ft

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table is closer to the surface, dewatering process becomes a need. Not only that the blasting agents and the

explosives also must be selected based on their ability to withstand wet conditions. Therefore water

resistant explosives should be used.

The following information are assumed in this design.

Shaft area = 5.2m2

Bit size of the drilling machine =38 mm

Advance of the shaft per round = 1m

Number of holes per round = 60

Powder factor for the blast = 3.1 kg /m3

Romanic Numbers half second delay

Arabic numbers m second delay

Figure 2.6 Main shaft & Auxiliary shaft blasting pattern

Dimension in ft

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Due to the presence of the river the water table might well be very close to the surface of the area.

Therefore the mine shaft conditions are ought to be wet. It is important to use gelatineous water-

resistant explosives. After drilling the holes must be blown out with compressed air. The premier

cartridge must be inserted first into the hole, the detonator pointing to the main charge. The final

trimmer holes should be drilled with a slightly smaller burden and loaded relatively lightly.

3.3 Shaft lining Since the area geology is to be assumed that of a strong self-supporting type we can use concrete

lining for the shaft. There is no seepage assumed because the shaft sinking location was selected

avoiding the buffer zone of the river. The shaft collar is developed to the bed rock and at the shaft

location it is assumed to be reaching the bed rock at a depth of 30 m. therefore the length of the

drill collar is considered as 30 m and the thickness of the concrete wall of the collar is set in three

different phases.

For the first 15 m from the surface the thickness of the shaft collar is set as 1 m

For the next 10 m it is set as 0.7 m.

From 25 m to 30 m the thickness is set as 0.5 m.

The winch is located at the same level where the top of the blind shaft lies, on top of the drill

collar, using a steel structure as shown in the figure below. A very accurate survey has to be done

to position the winch.

Figure 2.7 The winch and the steel structure

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3.4 Cross cut

In this particular design the cross cuts are designed to be arched shape and are to be driven with a dip of

0.005 towards the mouth. No extensive timbering is to be use. The cross cuts are designed to be of the

shortest length from the main addits to the ore bodies, so that the cost would be minimum.

The cross cuts are the developments used to reach the working areas of the mine. Therefore the

ventilation and the electricity supplies have to be transported through these. As shown in the

figure below the ventilation duct and the electricity lines are placed on to the left top corner of the

cross cut cross section. There should be a spacing in between those and the mine cars as well as

the men who are travelling in the cross cut. That is the reason why spacing is provided as shown

in the figure.

Figure 2.8 The cross cut with dimensions

Dimension in ft

Cross cut

Ventilation

Electrical Line Ore Transportation

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The following figure exhibits the drilling pattern as well as the blasting sequence at the same time.

To come to the conclusion regarding the number of drill holes and the advance of the shaft per

one cycle of the process a small calculation has been done as follows.

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Cross cut area = 4.1m2

Bit size of the drilling machine =38 mm

Advance of the shaft per round = 1m

Number of holes per round = 37

Powder factor for the blast = 2.5 kg /m3

Figure 2.9 The blasting pattern of the cross cut

Dimension in ft

Cut Hole (64 mm)

Drill Hole (32 mm)

0

0 0

0

ii

ii

ii

ii

Iv

iv iv

iv

Iv

vi

vi

iv

3 3 3

4

4

4

4

5

5

5

vi

6 6 6

7 7

8 9

9

9

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3.5 Mining method

3.5.1 Overhand cut and fill method

Overhand cut and fill mining is a highly selective open-stope mining method considered ideal for

steeply dipping high grade deposits found in weak host rock. Many variations of the general cut

and fill technique exist. Overhand cut and fill evolved from square set stull stoping to provide

stronger support. In this method, mining begins at the bottom of the ore body or block and

progresses upward.

During the mining sequence, the back of the excavation is to be temporarily supported using rock

bolts before the stope is back filled to form the floor of the next level of development. Backfill is

designed to provide mild excavation support as well as to provide a strong working floor for

personnel and equipment. Backfill selection is dependent on the quality of the host rock and the

size of equipment operating on top of the backfill.

Figure 2.10 The ore blocks created to be mined

In this particular design the height difference between the two levels is set to be about 40m. This

undercut will form the transport drift from which ore will be removed by rail cars. The winzes

were sunk from top level to bottom level in order to connect drives vertically. As we proceeds

upwards, first one or two benches are to be kept vacant and the backfilling from the next bench is

started. The reason for doing this is that we need that space to construct the haulage way and to

install machinery, such as winches, pumps, etc.

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Figure 2.11 The layout of the mine with cross cuts, winzes and shafts

Advantages and Disadvantages of this Mining Method

Advantages

High selectivity and low dilution ma achieved

Minimal development is required; low capital cost

Versatile for mining method; can follow irregular ore bodies

Flexible; mining method can be easily modified

Low equipment investment relative to other methods

Minimizes ground movement

Disadvantages

Cyclical ore production

Labour and skill intensive

Auxiliary shaft

Ore body

Drift

Winzes

Addit

Cross Cut

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Dangerous working conditions; work conducted on a top of freshly blasted rock

High degree of ground control required

Expensive and costly ventilation system

Need for backfill infrastructure (piping and paste plant)

Not suitable for low grade ore due to high mining cost

The extraction operation is done in the traditional way, i.e. drilling and blasting. The drift is to be

developed along the vein. The figure below illustrates the blasting sequence and the pattern.

Figure 2.12 The extraction procedure

The portion of the vein in a given cross sectional area is to be initially drilled, charged and

blasted. After that mucking has to be taken place. When that is finished the remaining portion of

the drifts cross section is drilled, charged and blasted to trim the edges in order to retain the shape

of the drift. After the execution of the whole operation of drilling, charging and blasting a total of

a 1 m advance is expected. At maximum of a two days duration the whole cycle is expected to be

finished. Hence the rate of advancement of a drift is 1 m per two days.

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3.6 Mine Layout

The below figure shows the whole layout of the mine discussed so far. The two blind shafts

connect to the auxiliary shaft through cross cuts. The auxiliary shaft is located at a position where

the optimum use of the natural ventilation flow can be utilized and the distances to the ore bodies

are minimized. Winzes and the drives are going along the ore body in different directions.

Figure 2.13 3D view of the mine lay out

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3.7 Underground transportation system

In this particular mine, rail tracks and locomotives are used as the underground transportation

system. The length of rails used in mines varies from 3 m to 7 m. The size chosen at a

particular mine often depends on the limitations imposed by the shaft and

transport arrangements. The aim should be to select the longest length of rail,

which reduces the number of joints needed when it is installed. This will also result

in a reduction in supply and maintenance costs. In this mine 7 m long rails are to be used.

The minimum weight of rail recommended for any new installation

underground in mines is 17.36 kg/m (35 lb/yd). Based on the expected production at the initial

stages of the mine these minimum weighted rails are to be used.

Figure 2.14 The dimensions of the rail cross section

In this particular design of the mine, the use of locomotives is omitted since the cut and fill mining

operation for such a narrow vein might not incur such a need.

The next most important part of the rail track is the use of sleepers. Sleeper spacing is the distance

between the centres of adjacent sleepers.

The following sleeper spacings have proven acceptable in practice and should not be exceeded

o locomotive track 840 mm;

o man riding rope-haulage track 840 mm;

o non-man riding rope-haulage track 1000 mmsince this design is for non-

manriding rope-haulage track, the spacing should be taken as 1000 mm.

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Mass of rail

(kg/m)

Maximum axle

load (tonnes)

Distance between

sleepers (mm)

Factor of

Safety

30.54 7.5 840 4.1

24.8 7.5 840 3.1

17.36 4 1000 2.5

Table 2-2 Factor of safety for the rails

The factor of safety for 17.36 kg/m rail is relatively low, and the maximum axle

loading of 4 tonnes is considered to be an abnormal load, which is carried only infrequently,

and not the regularly carried load. Fixed mine cars are used in this transport system. Their gauge

width is approximately 2 feet.

Considering the requirements MGC 1.7-6 mine cars can be recommended as more suitable for the

mineral transport operation (appendix iii).

Figure 2.15 The mine car with dimensions

The broken ore is directly transported to the mineral processing plant located at the point

specified. When the processing plant is to be located lot of factors have to be considered. The

accessibility of the location is the major point in a project like this which is situated in a hilly

terrain. Considering that and also the distance to be travelled from both the main addits, a location

for the processing plant is identified.

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3.8 The conveyor belt system

The ore has to be transported from the mine to the processing plant. In this design, there is a

conveyor belt system between the mine and the processing plant. This enables the continuous

operation of the mine. The guidelines for the design of the conveyor system are as follows.

Power system

Standardized three-phase squirrel-cage motors with star-delta start are preferable. The starting is

usually smooth. The belt speed is often controlled through an electronic frequency inverter.

Driving pulley

Here cylindrical conical shape driving pulleys are used since they will also produce a self-centring

effect on the running behavior of the belt. Clean, oil- and grease-free steel pulleys with a smooth,

almost polished surface (corresponding to a roughness Ra = 1.6 m) are in most cases sufficient to

ensure slip-free power transmission.

Pulley design

Since, conical cylindrical pulleys are used in the design, the following condition for the

admissible deflection y caused by the tension can be emphasized.

Figure 2.16 The deflection of the pulley

For Cylindrical-conical pulleys:

y (0.001 d) + 0.07 *mm+

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Pulley width

Pulleys and rollers should be wide enough so that the belt would make full contact with them over

its entire width, even when it is not positioned on the exact centre of the pulley.

Figure 3.17 Pulley width

Belt width bo Pulley width b

bo 100 mm b = bo + 20 mm

bo > 100 mm b = (1.08 bo) +

12 mm

Table 2-3 Pulley width recommendations

Multiple rollers are expected to be fixed on the path of the conveyor belt.

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3.8 Ventilation methods

3.8.1 Stope ventilation methods

The objective of any ventilation system is stage. First, the primary ventilation must course air

through the main airways to the immediate working area out by the working faces, thus making

fresh air available for face ventilation, and then return the contaminated air through return

(exhaust) airways to the surface. Second, the face ventilation system must be designed to

effectively utilize the available air in the immediate working area to sweep the working face, to

capture and remove dust, and to dilute and carry away gas, if any, emitted during mining

activities.

Without a properly designed ventilation system, an efficient production cycle would not be

possible. The system should provide the required air volumes and quality at reasonable pressure

losses, perform with minimum interference and cost to production, and do so in the most cost-

effective way possible. Furthermore, the primary ventilation system may be well designed, but if

the available air brought to the working area is not properly utilized for ventilating the faces

where most workers are located, the total system has failed.

Stope Ventilation

Separate ventilation shaft(in this design the auxiliary shaft) is used to the mine layout to throw the

air out from mine. Air is sucked from the mine main shaft down to lower level and air circulate

through all the working areas and used unclean air goes up through the winzes developed at the

ore body into ventilation shaft. Two or three Centrifugal fans can be installed on the surface at the

entrances of the two addits. Because of the fact that if one fan fails a specific amount of the total

air volume can be provide from rest. It is possible for the fan to be located directly on the top of

the auxiliary shaft. But usually air is turn through the right angle at the shaft collar and reach the

fan through an inlet drift and set of self-closing door. For Specification for centrifugal fan refer

appendix iii.

All stopes are ventilated by the fresh air flowing up from the lower level of the stope. There are

two raises constructed at each end of the stope. Air enters the raises at the bottom level and flows

up the raise and then enters the stope and exits through the other raise, depending on the pressure

difference. When stopes are mined out, the raises are closed, thus the movement of airflow was

affected. Similarly, when new raises are constructed for stoping, airflow is also affected.

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Therefore, there is a need to continuously monitor the movement of air especially in the stope

where miners are working. This study is conducted to assess the movement of air in the stope.

A mine ventilation booster fan is a primary fan that is in series with another primary fan. Without

the booster fan operaton, the total mine airflow will fall significantly. Booster fans are typically

installed as mine develops and primary fan no longer has the pressure or flow capability to service

increased resistance of operation. In this sense, booster fan operates as an integral and symbiotic

unit with original primary fan to provide sufficient total airflow for mine.

Figure 2.16 Stop ventilation plan

Mine booster fans are technically main fans which are installed underground to maintain required

airflow by overcoming the mine resistance. Booster mine ventilation fans can reduce pressure of

main fan and decrease system leakage and total required air power.

A mines booster fan is an underground ventilation device installed in the main airstream (intake or

return) to handle quantity of air circulated by one or more working districts. It is installed to

operate in series with a main fan and boost air pressure of ventilation air passing through it.

Specifications for the booster fans are given in appendix iv.

Air input

Air output

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3. REFERENCES

http://www.slideshare.net/hzharraz/lecture-1-9998430

o Accessed on 24-11-2014

Evaluation of Overhand Cut and Fill Mining Method used in Bogala Graphite Mines,

SriLanka; P G R Dharmaratne, P V A Hemalal and M C Hettiwatte

o Accessed on 22-11-2014