underground mine planning and design
TRANSCRIPT
Unit Code: 307320 Mark /Grade : 81
Unit Name: Underground Mine Planning and Design 601
Unit Coordinator: Dr Jayantha Bhattacharya
Project 1: Mine Access Selection
I declare that this assessment item is my own work, except where
acknowledged, and it has not been submitted for academic credit
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be found at:
http://students.curtin.edu.au/administration/responsibilities.cfm).
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 𝐿𝑒𝑛𝑔𝑡2 + 𝑉𝑒𝑟𝑡𝑖𝑐𝑎𝑙 𝑙𝑒𝑛𝑔𝑡2
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