Unit Code: 307320 Mark /Grade : 81

Unit Name: Underground Mine Planning and Design 601

Unit Coordinator: Dr Jayantha Bhattacharya

Project 1: Mine Access Selection

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Name: TITTU BABU - 16320587

Date: 22/10/2013.

Contents

1. EXECUTIVE SUMMARY ................................................................................................................ 2

2. INTRODUCTION ............................................................................................................................. 3

2.1. Scope and methodology: ......................................................................................................... 3

2.2. Background: .............................................................................................................................. 3

2.3. Objectives: ................................................................................................................................. 4

3. GEOLOGY ........................................................................................................................................ 4

4. VENTILATION ...................................................................................................................................... 8

5. MINING METHOD AND SELECTION .......................................................................................... 9

6.ASSUMPTIONS ................................................................................................................................... 10

7. Decline ............................................................................................................................................. 11

7.1Equipment Selection ................................................................................................................ 11

7.2TRUCKS ................................................................................................................................. 11

7.2.1CAT AD 30 ............................................................................................................................. 11

7.2.2CAT AD45 .............................................................................................................................. 12

7.2.3CAT AD 60 ............................................................................................................................. 12

7.3Length of Decline .................................................................................................................. 14

7.4Productivity ............................................................................................................................ 15

7.5Cycle Time ............................................................................................................................. 18

7.6Annual operating cost .............................................................................................................. 22

7.7Total cost of trucks ................................................................................................................... 22

7.8Load Haul Dump ....................................................................................................................... 23

7.9JUMBO ....................................................................................................................................... 23

8. MAN POWER REQUIREMENTS .......................................................................................................... 23

9. PRODUCTION SCHEDULING .................................................................................................... 24

10. Risk assessment .......................................................................................................................... 26

10.1MAJOR HAZARDS ................................................................................................................. 27

10.1.1DECLINE ............................................................................................................................... 27

10.1.2Shaft .................................................................................................................................... 28

11. Occupational health and safety of employees ........................................................................ 29

12. Conclusion .................................................................................................................................... 30

13. Recommendations .......................................................................................................................... 31

14. References ................................................................................................................................... 32

15. Appendix ......................................................................................................................................... 33

1. EXECUTIVE SUMMARY

In this hard rock metalliferous project, I critically analysed the various design and

development features of the major access system to the underground mine. The

Kalgoorlie Consolidated Gold Mine (KCGM) is developing a mine to extract gold ore

deposit from a proposed area which is near Kalgoorlie Western Australia. The

project is mainly used to develop a suitable access system to the underground mine.

Based on the studies performed on methodology, development and cost of

operations different access systems are studied. Decline was found to be the most

suitable of all and was selected as the most appropriate access system. A total of 4

trucks were selected in total during the full phase of production. A decline gradient of

1:6 was selected with a total development cost of 1449532.8AU$. A decline cross

section of 6m*6m was selected corresponding to the truck height.

2. INTRODUCTION

An Underground mining system planned to extract a deep lying ore body of gold-

copper. The peculiarity of underground mines is that, it’s been used for the extraction

of deep lying ore bodies at a greater risk. A proper access system is one of the major

component regarding the underground system. A well-developed safe access

system is important for the transportation of workforce, equipment’s and the ore from

the underground mines. Hence a good access system is very important for the

proper running of underground process.

2.1. Scope and methodology:

This report is used for the development of a suitable access system for an

underground mine which planning to extract deep lying ore body. Using a detailed

study regarding the strategy, methodology and scheduling the decline, adit and shaft

systems where been compared. Each access systems is been considered as the

best in certain cases, so studying the depth and geology of the ore body the most

suitable access system is been selected

2.2. Background:

A Gold-copper deposit is been planned to extract by the KCGM using the room and

pillar method. The owners have already adopted a design for the extraction of ore

using a drill and blast, mechanised room and pillar stoping system. The ore body is

been located about 100 km from a small country town called Kalgoorlie which is

located in Western Australia. An average temperature range of 10◦C-36◦C is

obtained during the time of summer whereas 4◦C-25◦C is delivered during the winter

period. Town is well developed with all facilities like road, rail and an airport for small

aircrafts. An electrical power line of 66kV passes adjacent to the proposed mine area

from the State Grid to the electrical power station in the town. Disturbance to the

surface must be minimal due to the severe access constraints as well as due to the

heritage and Native title considerations.

2.3. Objectives:

The main objectives of the project are.

Underground metalliferous mining development and stoping systems and key

factors used in their selection and performance.

Equipment and personal requirements.

Core risk identification and mitigation.

Technical and economic aspects.

Environmental and OHS considerations.

Technological trends.

Legal and statutory requirements.

3. GEOLOGY

The Au-Cu deposit is been found after extensive regional and local geophysical

studies and also by limited core drilling technique. With the help of other geological

studies it’s been found that there are 3 mineable ore bodies existing in the proposed

area. All the ore bodies are at a depth of 80-120 m below the surface which is found

to be known after further drilling.

Each three ore bodies found to have a deposit of about 3.7Mt in NE, 3.0Mt in W and

3.5Mt in the S ore body. The ore bodies are found to be sandwiched in between

dolomite and limestone and is been layered in structure. The shape of three ore

bodies found to be flat with a thickness ranging from 2-25 m. A total mineable

resource of 10.2Mt will be obtained from the three ore bodies. Most of the ore bodies

is been hosted by dolomite with a layered limestone as the footwall. It has got a

uniaxial compressive strength of 100Mpa and Young’s Modulus of 12MPa. The

average solid density of ore body is found to be 3.5 t/m3 and for the roof and floor

rocks it’s found to be 2.7 t/m3.

The company has proposed to use a stoping method of room and pillar for the

extraction of ore body due to the severe access constraints existing in the area.

Using the UBC method it’s found that the Room and Pillar comes second best

behind the Open pit mining method, since the area is severely constraint and due to

the heritage and Native Title considerations it’s been opted to take room and pillar

method.

Figure 1 room and pillar mining

An output of 2500 tonnes/day is estimated to be delivered from a relatively small

workforce. With this expected output per day a total of 0.75 Mt/annum is produced.

The production is to start from the NE ore body as proposed by the company. After

full development works towards the NE ore body is done and production started then

decline towards W ore body begins. Later on the S ore body is been extracted. No

more than one active area should be exist at a single time hence the three ore

bodies is been extracted at different time.

The layout of the ore bodies is shown below.

Figure 2 layout of ore body

The geology of the three ore bodies shows that it’s been deposited in between

dolomite and limestone.

Figure 3 geology of NE ore body

Figure 4: geology of W ore body

4. VENTILATION

Ventilation in underground mines is very important for the proper working environment and

for the occupational health and safety for the employees. Underground mines are usually

confined in space and is with too much dust, diesel fumes and other dust particles. So a

proper working environment is not possible in underground mines without the proper

ventilation system. With the help of efficient and properly working ventilation system it is

able to maintain the level of temperature, humidity and air velocity to a much suitable

condition for the working. So a correct and proper ventilation is important for the

underground mines. The basis of ventilation is the primary ventilator system which consist of

a primary fan which helps in boosting the air speed through the depth of underground mines.

Ventilation also consists of auxiliary fans which helps in the supplying of air through the

different levels of underground mines. Correct selection of auxiliary fans depends upon the

diameter of the duct, amount of energy required etc. ventilation system is of different type’s

series and parallel ventilation. An effective ventilation happens by the integrated working of

both the primary and secondary ventilator systems.

Figure 5 geology of s ore body

Figure 6: parallel ventilation Figure 7: series ventilation

In this case of underground mine, series ventilation is been used. For good and higher

productive mines series ventilation is much suitable. The size of ventilation duct is made to

be 6m*6m which uses the fresh air effectively. Series ventilation is much simpler and easier

in operating. It also reduces the need of ventilation officer. Series ventilation helps in

reducing the number of ventilation control devices such as regulators. Series ventilation

circuit also offers less susceptibility to leakage and recirculation.

5. MINING METHOD AND SELECTION

As per geology and ore reserves of our deposit, the deposit consists of three major

ore bodies NE(3.7 Mt),W(3 Mt) and S(3.5 Mt) at a depth of 80m to 120 m below

surface. Moreover all ore bodies are located in a layered structure between

sedimentary layers of dolomite and limestone. And thickness of the ore is

intermediate and shape is irregular. According the above information we found out

the best mining system using UBC method. The below table shows the summery of

the UBC method.

Mining method General shape

Ore thickness

Ore plung

Grade distribution

Total

irregular intermediate Flat Gradational

Open pit mining 3 3 3 3 12

Block caving 0 0 3 2 5

Sublevel stopping 1 2 2 3 8

Sublevel caving 1 0 1 2 4

Longwall mining -49 0 4 2 -43

Room and pillar mining 2 2 4 3 11

Shrinkage mining 1 2 2 2 7

Cut and fill stopping 2 4 0 3 9

Top slicing 0 0 4 2 6

Square set stopping 3 3 2 2 10

Table 1 UBC METHOD

As per UBC method we selected four suitable mining method, in that open pit mining

method is the best option for mining. But due to heritage , native and geotechnical

issues we are not using open pit, so we choose the second option room and pillar

mining as the mining method for this underground project. The project summary also

recommends the same mining method for this project. Based on the values other

alternative mining methods are cut and fill and square set stoping.

The stoping design will be the nominal room and pillar with a stope size 8m wide by

4m to 8m high. Remaining pillars will be 6mx 6m.In thicker parts of the ore, initial

stope will be extended to a final height of maximum 25m by horizontal and vertical

benching.

6.ASSUMPTIONS

Ventilation duct diameter =1.4 m

Space to roof clearance=0.1 m

Ventilation duct clearance = 0.3 m

Annual production days =360 days

Working hours per day = 2*12 hour shift

Mine stockpile located 1km from surface

Mine waste dump is located 300 m from portal

Transport efficiency factor is considered as 90%

Truck availability factor is considered as 90%

7. Decline

7.1Equipment Selection

In the case of mines equipment selection is a critical part. The Equipment in a mine

includes Trucks, LHDs, Jumbos, Bolters, lights vehicles etc. These equipment are

used for a unique purpose. The equipment listed above are selected based on

certain criteria’s. These criteria’s are explained below.

7.2TRUCKS

7.2.1CAT AD 30

Figure 8: CAT AD 30

Payload capacity of 30 tonnes

Designed for high production

Low cost per ton

Suitable for smaller underground operations

Rugged construction

Easy maintenance

Long life

Low operating costs

7.2.2CAT AD45

Figure 9: AD 45

Payload capacity of 45 tonnes

Simplified maintenance

Designed for high production

Excellent fuel efficiency

Lowered noise

7.2.3CAT AD 60

Figure 10: CAT AD60

Payload capacity of 60 tonnes

Deliver excellent fuel efficiency

Lower emissions

Reduced noise

Lower operating cost

Cross-sectional Dimensions

It is assumed that,

Ventilation duct diameter =1.3 m

Space to roof clearance=0.3 m

Ventilation duct clearance = 0.5 m

Figure 11: Dimension of trucks

In order to find out the cross sectional size of decline for each type of the truck the

total height has to be found which is given by the formula

Total height = Height to top of load + vent duct clearance + vent duct diameter + duct

to roof

The cross sectional size of each trucks are calculated and summarized in table 1.

Truck Selection AD 30 AD 45 AD 60

Cross section 5.5m x 5.5m 6m x 6m 6m x 6m

Height to top of load(m)

3.04 3.705 3.848

Vent duct clearance 0.5 0.5 0.5

Vent duct diameter 1.3 1.3 1.3

Duct to roof 0.3 0.3 0.3

Table 2 cross sectional size

7.3Length of Decline

Length of decline is calculated using Pythagoras theorem.

Decline length = 𝐻𝑜𝑟𝑖𝑧𝑜𝑛𝑡𝑎𝑙 𝐿𝑒𝑛𝑔𝑡𝑕2 + 𝑉𝑒𝑟𝑡𝑖𝑐𝑎𝑙 𝑙𝑒𝑛𝑔𝑡𝑕2

Vertical distance is the depth of the decline which is given as 120 m and the

horizontal distance varies according to the different gradients.

For gradient 1:6, Decline length = 7202 + 1202= 730 m

For gradient 1:7, Decline length = 8402 + 1202 = 849 m

For gradient 1:8, Decline length = 9602 + 1202= 967 m

1.4 Development Cost

Development cost depends on the cross sectional sizes. For different cross sectional

sizes development cost will be different. Development cost for different cross-

sections is summarised in table 3.

Length of Decline

Depth of

Decline

Profile 5.5mx5.5m 6mx6m

Cost data($

𝒎) 7000 8000

Table 3development cost per unit meter

Decline development cost ($) =Decline length (m) × Development cost ($/m).

Using this formula development cost is calculated for each cross-sections, for three

different gradients which is summarised in table 4.

Total Development

Cost($)

Cross section

selected

gradient 1:6 gradient 1:7 gradient

1:8

AD 30 5.5m x 5.5m 5110000 5943000 6769000

AD 45B 6m x 6m 5840000 6776000 7736000

AD 60 6m x 6m 5840000 6776000 7736000

Table 4 total development cost for different gradient

From the table 4 it is clear that AD 30 which has the smallest cross-sectional size

has the lowest development cost. The gradient option 1:6 is more suitable as it has

the lowest development cost because of its shorter length.

7.4Productivity

Speed of Truck

Speed of the truck has to be calculated in order to find the cycle time of the

trucks.Rimpull-speed-gradeability chart is used to find the speed of the truck.This

particular chart is obtained from the CAT underground performance

handbook.Rimpull-Speed-Gradeability chart makes use of the total resistance which

the truck experiences.

Total resistance is the sum of rolling resistance and grade resistance. By referring

the Caterpillar performance handbook it is been found that, Rolling resistance is

taken as 2% for underground operations. Grade is defined as the force that must be

overcome in order to move a truck uphill. It is the ratio between the vertical rise and

the horizontal distance in which the rise occurs.

Loaded and Empty speed for truck AD30

Figure 12 speed resistance graph for AD 30

Figure shows the Rimpull-Speed-Gradeability chart for CAT AD30 truck for three

different gradients. Using the total resistance, a horizontal line plotted from the total

resistance point. This line intersects the curve with highest achievable gear. A

vertical line is dropped from this intersecting point and it meets the X-axis which

gives the speed of the truck.

Cross section

selected

gradient

1:6

gradient

1:7

gradient 1:8

Resistance 5.5m x 5.5m 0.16 0.14 0.125

Total resistance 0.18 0.162 0.145

AD 30(loaded) 7.6 8.5 9

Empty 17.2 17.2 17.2

Table 5 truck speed

The same procedure is followed for the all the trucks and the speed is calculated.

Loaded and Empty speed for truck AD45

Figure 13 speed resistance graph for AD

Cross section

selected

gradient 1:6 gradient

1:7

gradient 1:8

Resistance 6m x 6m 0.16 0.14 0.125

Total resistance

AD 45(loaded) 0.18 0.16 0.145

7.9 8.9 9.4

Empty 22.5 22.5 22.5

Table 6 truck speed of AD 45

Loaded and Empty speed for truck AD30

Figure 14 speed resistance graph for AD 60

Cross section

selected

gradient

1:6

gradient

1:7

gradient

1:8

Resistance 6m x 6m 0.16 0.14 0.125

Total resistance

0.18 0.162 0.145

8.6 10 11

AD 60(loaded)

Empty 25.2 25.2 25.2

Table 7 truck speed for AD60

7.5Cycle Time

Total cycle time includes hauling time, loading time and dumping time. Table below

shows the given loading and hauling time for each truck types.

Truck Model AD 30 AD 45B AD 60

Total Tonnage 30 45 60

Loading Time(min) 1.5 2 3

Dumping Time(min) 0.5 0.75 1.25

Table 8 loading and haulage time for each truck

Grade Gradient

1:6

Gradient

1:7

Gradient

1:8

Width 720 840 960

Depth 120 120 120

Decline metres 730 849 967

Table 9 decline dimension for each truck

Hauling time = Loaded time + empty time

Loaded time = Ore to portal + portal to pit surface + Pit to stockpile

Empty time = Stock pile to pit + Pit to portal + Portal to ore

Formulas for calculating Loaded time of the truck

Ore to portal=𝐿𝑒𝑛𝑔𝑡 𝑕 𝑜𝑓 𝑑𝑒𝑐𝑙𝑖𝑛𝑒

𝑆𝑝𝑒𝑒𝑑 𝑜𝑓 𝑎 𝑙𝑜𝑎𝑑𝑒𝑑 𝑡𝑟𝑢𝑐𝑘

Pit surface to Stock pile=𝐿𝑒𝑛𝑔𝑡 𝑕 𝑜𝑓 𝑝𝑖𝑡 𝑠𝑢𝑟𝑓𝑎𝑐𝑒 𝑡𝑜 𝑠𝑡𝑜𝑐𝑘𝑝𝑖𝑙𝑒

𝑆𝑝𝑒𝑒𝑑 𝑜𝑓 𝑎 𝑙𝑜𝑎𝑑𝑒𝑑 𝑡𝑟𝑢𝑐𝑘;

Formulas for calculating the empty time of the truck

Stockpile to pit surface=𝐿𝑒𝑛𝑔𝑡 𝑕 𝑜𝑓 𝑝𝑖𝑡 𝑠𝑢𝑟𝑓𝑎𝑐𝑒 𝑡𝑜 𝑠𝑡𝑜𝑐𝑘𝑝𝑖𝑙𝑒

𝑆𝑝𝑒𝑒𝑑 𝑜𝑓 𝑎 𝑒𝑚𝑝𝑡𝑦 𝑡𝑟𝑢𝑐𝑘

Portal to ore=𝐿𝑒𝑛𝑔𝑡 𝑕 𝑜𝑓 𝑑𝑒𝑐𝑙𝑖𝑛𝑒

𝑆𝑝𝑒𝑒𝑑 𝑜𝑓 𝑎 𝑒𝑚𝑝𝑡𝑦 𝑡𝑟𝑢𝑐𝑘

Before finding the cycle time for each truck type. Another assumption is taken in to

account. The timings of various jumbo operations are assumed as following.

Drilling time – 2.5 hours

Charging and firing – 1.5 hours

Re-entry – 30 min

Wash Down – 20 min

So in total it will take around 5 hours to complete all this processes. With 24 hours as

operating hours it is able to complete 4 cuts using a jumbo per day.1 Jumbo cut is

assumed to advance 3.5m.As a result it will be possible to advance nearly 21 m per

day. So initially there is no need of too many trucks. Table below shows the cycle

time and the number of trucks required in the initial stage to the final stage for each

truck model.

Dev

m

Total

Travel m

Loaded

travel(min)

Empty travel

(min)

Total Time

AD30

No of AD 30

Trucks

14 314 2.48 1.10 5.02 2

28 328 2.59 1.14 5.16 2

42 342 2.70 1.19 5.30 2

56 356 2.81 1.24 5.45 2

70 370 2.92 1.29 5.59 2

84 384 3.03 1.34 5.73 2

924 1224 9.66 4.27 14.65 4

938 1238 9.77 4.32 14.79 4

952 1252 9.88 4.37 14.93 4

966 1266 9.99 4.42 15.06 4

980 1280 10.11 4.47 15.20 4

Table 10 loaded and empty traveling time for AD30

Dev

m

Total

Tavel

m

Loaded

travel(min)

Empty

travel

(min)

Total

Time

AD45

No of AD

45

Trucks

14 314 2.22 0.84 5.80 1

28 328 2.32 0.87 5.94 1

42 342 2.41 0.91 6.08 1

56 356 2.51 0.95 6.21 1

70 370 2.61 0.99 6.35 1

84 384 2.71 1.02 6.48 2

924 1224 8.64 3.26 14.65 3

938 1238 8.74 3.30 14.79 3

952 1252 8.84 3.34 14.93 3

966 1266 8.94 3.38 15.06 3

980 1280 9.04 3.41 15.20 3

Table 11loaded and empty traveling time for AD45

Dev

m

Total

Tavel

m

Loaded

travel(min)

Empty

travel

(min)

Total

Time

AD60

No of

AD 60

Trucks

14 314 2.09 0.75 7.09 1

28 328 2.19 0.78 7.22 1

42 342 2.28 0.81 7.34 1

56 356 2.37 0.85 7.47 1

70 370 2.47 0.88 7.60 1

84 384 2.56 0.91 7.72 1

910 1210 8.07 2.88 15.20 2

924 1224 8.16 2.91 15.32 2

938 1238 8.25 2.95 15.45 2

952 1252 8.35 2.98 15.58 2

966 1266 8.44 3.01 15.70 2

980 1280 8.53 3.05 15.83 2

Table 12loaded and empty traveling time for AD60

Number of trucks required for each gradient for each model is shown in the table

below.

Number of trucks

gradient 1:6

gradient 1:7

Gradient 1:8

AD30 2 3 4

AD45 2 2 3

AD60 2 2 2

Table 13truck required for each gradient6.5Truck Capital cost

TYPES OF TRUCKS CATIPAL COST PER

UNIT($)

AD30 982080

AD45B 998910

AD60B 1970100

Table 14truck capital cost for each model

Using the number of trucks for each gradient for each model the capital cost for the

trucks are obtained.

Total capital cost for

trucks in ($)

Gradient 1:6 Gradient

1:7

Gradient

1:8

AD30 1964160 2946240 3928320

AD45B 1997820 1997820 2996730

AD60B 3940200 3940200 3940200

Table 15: Capital cost for trucks.

7.6Annual operating cost

Given information

AD30 AD45B AD60B

Operation

cost/Hour in $

107.77 137.79 226.98

Driver Salary/hour

in $

60

Table 16: annual operating cost

Annual operating cost is calculated using the formula

Annual operating cost = 𝑜𝑝𝑒𝑟𝑎𝑡𝑖𝑛𝑔 𝑐𝑜𝑠𝑡 𝑝𝑒𝑟 𝑕𝑜𝑢𝑟 × 𝑛𝑢𝑚𝑏𝑒𝑟 𝑜𝑓 𝑕𝑜𝑢𝑟𝑠 𝑝𝑒𝑟 𝑦𝑒𝑎𝑟 ×

𝑜𝑝𝑒𝑟𝑎𝑡𝑜𝑟 𝑠𝑎𝑙𝑎𝑟𝑦 𝑝𝑒𝑟 𝑦𝑒𝑎𝑟

The annual operating cost for each model is summarized in the table below.

Annual Operating

Cost

AD30 AD45B AD60B

Operation cost/Hour 107.77 137.79 226.98

No of hours/Year 8640 8640 8640

Total Operating cost

without driver salary

931132.8 1190505.6 1961107.2

Total Operating cost

with driver salary

1449532.8 1708905.6 2479507.2

Table 17: total operating cost.

7.7Total cost of trucks

Total cost is calculated using the formula

Total Cost = Decline Development cost + Total Capital Cost + Total operating Cost

Table below shows the total cost for each gradients.

total cost in $ Gradient 1:6 Gradient 1:7 Gradient 1:8

AD30 8523692.8 10338772.8 12146853

AD45 9546725.6 10223352.8 12441636

AD60 12259707.2 12425105.6 14155707

Table 18: total cost for each gradient.

From the table it is clear that Truck AD30 with a gradient of 1:6 gives the least capital

cost.

7.8Load Haul Dump

LHD unit having a bucket capacity 3.82 m3 preferred and it cost nearly 5 million

dollars. LH 307 is the model that matches with these specifications with a bucket

capacity of 13.7 t.

Two such units are required to satisfy the production needs.

Capital cost for the LHDs = 0.755 × 2 = 1.51 M$

7.9JUMBO

In order to drill a 5.5m × 5.5m cross sectional face a 2 boom jumbo with a boom span of 6.34 m × 8.84 m is preferred. Only one unit of such jumbo is required.

Capital cost for the jumbo = 0.91 M$

8. MAN POWER REQUIREMENTS

For the proper working of any operation there must be good workforce requirement. During the early development of decline the number of man power required is comparatively less. But at the final period during the full production a larger number of work force is needed.

Type of operation No.of employees

Shift boss 1

Charge up operator 2

Loader operator 2 (needed 1 at the beginning)

Truck driver 4 (only needed 2 at beginning)

Jumbo operator 1

Jumbo offside 1

Long hole operator 2 (no need at the development of decline)

Maintenance person 4

Underground electrician 1

Total employees for a single shift 18

Table 19: employee details in each shift.

A total crew of 18 members is being considered for a single shift. For the initial stages of decline development the number of employees required are comparatively less during the period of full development. A 2:1 roster is being prescribed for the proper working of the underground mine with a total of 3*18 employees is been selected. A total of 2 shifts are considered for a single day with 12 hour per shift.

9. PRODUCTION SCHEDULING

Scheduling can be defined as the distribution of available resources in terms of years

or months to meet objective of the company. According to RG Schroeder (2000) it is

the final and most constrained decision in the hierarchy of mine planning decisions.

There is no particular method for production scheduling it varies time to time and

place to place. So a scheduling method for one mining company may not necessarily

satisfy the requirements of another mine due to several reasons. Thus, the choice of

a scheduling method/software depends on nature of mining method we are using,

production requirement mill capacity and many more.

First logbook file is created which shows the number of operational shifts per day,

days per week and the number of years the production will stay in effect, in addition

to that an equipment list will prepare for knowing what machinery will operate

throughout the scheduling , their individual productivities and their cycle time.

Mostly we are using this two lists for scheduling the production of the mine. Then we

can use excel or other suitable scheduling software’s to do the rest of the scheduling

part .in this project we used smart draw software. The result we are getting after

putting all the given dates is a summary of the logical or scheduling.

After scheduling numerous reports of production, timing and activity, equipment

sequence etc. can be generated, as well as bar charts and coloured plots shaded by

period can be generated.

According to our project background our total mineable reserves is 10.2Mt.and it is

divided into three sections NE(3.7Mt), W(3Mt), S(3.5Mt).our daily production is set to

be 2500tonnes per day and annual production is 0.75Mt.depends on this data we got

a mine life of 13.6 years. Inspite of this we expected to start at first of Jan 2015 and

ends at 13th of march 2029.During this long years the first one and a half years will

be development work and end of the first year we will start the production in the NE

ore body. in NE ore body we divided into 4 panels and each panel takes 1 year for

completing the production and a total of five years for the NE ore body. Apparently

west ore body will start mining at the end of six year it will took 4 years to complete

and it consists of 2 panels and 2 years for each panel. And will start the south ore

body at the end of 10th year and it took 4.5 years to complete consists of 2 panel 2.2

years for each panel. Overall 14.5 years to complete this project. The below Gantt

chart shows the scheduling of the project

Figure 15 production scheduling

10. Risk assessment

Risk assessment is distinct as the result of consequence (hazards) and Likelihood

(probability of occurrence of hazard).

Risk assessment for hazards

Hazards consequence Likelihood Severity

Decline

Seismic activity

displacement of rock mass around the excavation

2 8 Moderate

Ground Water conditions

increase the development cost and time

1 3

Negligible

Blast Damage

over-break or under-break

3 5 Moderate

Decline in flexibility

increase the development cost and time

2 3 Negligible

Safety hazard

Under ground operations are often face vehicle crash which causes severe injuries to the workers

Human errors towards these accidents.

4 5 Extensive

Shaft

Piston effect

It creates a huge health a safety risk and can endanger the people working in and around the area.

2 6 Moderate

Electrical supply

Shaft operations really require high voltage power supply it increase the maintenance cost.

2 3 Negligible

Risk assessment= consequence* Likelihood (probability)

10.1MAJOR HAZARDS

10.1.1DECLINE

10.1.1.1Seismic activity

In mining the underground excavation will always give rise to seismicity which will

leads to the displacement of rock mass around the excavation. The after effect of

this is wall failure of decline which will badly affect the safe working condition.

Besides this the development cost of the project will also turns huge.

10.1.1.2Ground water conditions

Another major problem for decline development could be the ground water in

underground excavation. For underground workers including engineers the

underground water management will be a challenge. It will increase the development

time and cost. By installing high power motors, a well planned ground water system

can be employed. It can also be managed by proper grouting.

10.1.1.3Blast damage

Utmost care should be given while blasting. Poor blasting will cause under-break or

over-break in declines. An under-break will bring hangings in the decline, in that case

secondary blasting will require. This will increase development time and cost. The

following are the major reasons behind poor blasting,

Inexperienced workers

Selecting inappropriate explosives

Shaft inflexibility

The failure of bore holes can cause mine closure for it is the only access to the underground workings.

3 3 Negligible

Geotechnical issues

Geotechnical conditions will vary at increasing depth. Any failure in shaft will cause irretrievable lose to the project

4 2 Negligible

Table 20risk assessment

10.1.1.4Decline flexibility

Always the plans for Decline developments are not flexible. If any unfavourable

geotechnical conditions arise the layout of decline cannot be changed. Also it is not

possible to change the equipments selected such as truck and LHDs for different

layouts as it makes additional maintenance cost.

10.1.1.5Safety Hazard

Workers often face vehicle crash in underground operations which may cause

severe injuries to them. These accidents are mostly caused due to improper lightings

and lack of safety signs. One of the main issues is the mistakes due to unskilled

workers.

10.1.2Shaft

The technical risks which were identified and assessed are:

10.1.2.1Piston effect

There are many risk associated with shaft one of the main risk is the piston effect it

happens when the raise bore, slip and line option is adapted for shaft sinking a plug

of rock produced when the slipping is carried out. This causes an air blast will form at

the bottom of the blast and top of the shaft at the surface. This will make a huge

health and safety risk issues and it cause danger to people working in that area and

around that area.

10.1.2.2Electrical supply

A high voltage power supply is required for shaft operations. Unlike decline operation

a constant level of power supply should be maintained for which a high power

transformer is required. The maintenance cost will automatically increase by this.

Power supply hazards will always exist. Due to ground water conditions power

cables should be properly insulated.

10.1.2.3Shaft inflexibility

If there arise any failure of bore holes then it will cause mine closure because the

only access to the underground workings is this bore holes. Until the problem is fixed

the operations will be idle. So it is important that a considerable amount should be

invested in shaft maintenance.

10.1.2.4Geotechnical issues

A good geotechnical requirement is required for shaft sinking for its smooth

operations. At an increasing depth the geotechnical conditions will vary. If there

occurs any failure in shaft then the project will face irretrievable loss.

11. Occupational health and safety of employees

OHS of employees includes identification main mining hazards like

slope stability,

rock falls,

gas outbursts,

loss of ventilation

Safety duties of mine operators

identification of hazards and assessment and of risks

who may enter a mine

alcohol and drugs

employee fatigue

health surveillance

Consultation and information

consultation with employees and health and safety representation

information , instruction and training

information to visitors

information to job applicants

Duties of employees

use the appropriate PPE

Participate in the testing of the emergency plan

Follow the emergency plan when it is activated

Figure 16 OHS PLAN

12. Conclusion

The overall cost for decline was found to be 8523692.8$ . By taking into

consideration the overall cost, geotechnical issues with shaft sinking and according

to risk assessment number. Decline is the most suitable method for this project with

greater flexibility of mine expansion for variation in geotechnical conditions and lower

risks involved.

Truck AD 30 utilizes its full capacity and requires only 2 trucks in the initial stage and

using 4 trucks at the end to meet the annual production requirements at gradient

1:6.And has the total capital cost of 1964160$ at gradient 1:6 which is less when

compared to the other two trucks .So while considering the overall decline

development cost truck, capital cost and operating cost, AD30 will be the right choice

for the operation at gradient 1:6 as only 4 trucks is required to meet the actual

demand. A cross section of 6m×6m has to be selected based on the total height

calculated which has the unit development cost of 1449532.8AU$.

13. Recommendations

At the end of our detailed study of the project we felt that we can improve many

areas were we did already so we are taking this section to tell some

recommendation that we felt after this project. We should implement some changes

in mine planning and implement more underground activities at the site I hope the

information and knowledge from initial development helps to achieve this. Here we

are using room and pillar method throughout the operation I would like suggest that

in harder zones we should use long wall stopping method. We should install proper

ground supports in weaker areas. Other main drawbacks we saw in this project is we

didn’t consider the time factor. Time to time their will be a change in economic data it

changes the total development cost, commodity cost and overall cost we suggest

that we should consider the time factor. In addition to that more safety measures we

should consider for better and safe environment for the workers. One main reason

for accidents is lack of communication we should communicate with our co-workers

always during the work and outside. Other few things we would like to recommend is

installing the road signals in mines, safe storage system for fuels and explosive,

proper water handling system and proper maintenance for all machines and maintain

a chart for maintenance.

14. References

Visser D, Murray & Roberts Shaft sinking method based on the town lands

ore replacement project- raise boring shaft sinking and mining contractors

conference 2009

K matusi, Underground mining transportation systems K ,Kyushu University,

Fukuoka, Japan http://www.eolss.net/sample-chapters/c05/e6-37-06-07.pdf

Web link: FLSmidth product index: mine shaft systems

http://www.flsmidth.com/en-

US/Products/Product+Index/All+Products/Underground+Mining/Cages/Cages

Matsui. K. Underground mining transportation systems. 2011. Kyushu

University, Fukuoka, Japan. Civil Engineering – Vol. II.

Brazil. M, Grossman. N. C, Wormail. D. H, Rubinstein and Thomas. D.A.

Decline design in underground mines using constrained path optimisation.

2008. 561-578.

Matunhire. I. M. Design of Mine Shafts. Department of Mining Engineering,

2007. University of Pretoria, Pretoria, South Africa.

Caterpillar Performance Handbook. 2011. Underground Mining Truck AD30.

www.cat.com

Caterpillar Performance Handbook. 2011. Underground Mining Truck AD45B.

www.cat.com

Caterpillar Performance Handbook. 2011. Underground Mining Truck AD60B.

www.cat.com

15. Appendix