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REVIEW AND IMPROVEMENT OF UNDERGROUND VENTILATION
CONDITION AT 3800M REDUCED LEVEL WITH REFERENCE TO
DEEP CENTRAL MINE VENTILATION SYSTEM AT BULYANHULU
GOLD MINE LIMITED (BGML)
ALLEN EMMANUELLY DIOCLES
Reg. No. 428 MID10
Ordinary Diploma in Mining Engineering
Mineral Resources Institute
July 2012
ii
REVIEW AND IMPROVEMENT OF UNDERGROUND VENTILATION
CONDITION AT 3800M REDUCED LEVEL WITH REFERENCE TO
DEEP CENTRAL MINE VENTILATION SYSTEM AT BULYANHULU
GOLD MINE LIMITED (BGML)
By
Allen Emmanuelly Diocles
Technician Certificate in Mining Engineering
A project work Submitted in Partial Fulfillment of the Requirement for the Ordinary
Diploma in Mining Engineering of the Mineral Resources Institute
Mineral Resources Institute
April 2013
i
CERTIFICATION
This is to certify that the project have read and hereby recommended for acceptance by the
Mineral Resources Institute a project work entitled: Review and Improvement of
Underground Ventilation Condition at 3800M reduced level with Reference to Deep
Central Mine Ventilation System at Bulyanhulu Gold Mine Limited (BGML), submitted for
the award of Ordinary Diploma in Mining Engineering of the Mineral Resources Institute. To the
best of my knowledge, the matter embodied in the project has not been submitted to any other
University/Institute for the award of any Degree, Diploma or Full Technician.
………………………………………
Prof/Dr/Mr. /Ms.
(Supervisor)
Date…………………………….
………………………………..
Prof/Dr/Mr./Ms.
(Supervisor)
Date...........................................
.........................................................
Prof/Dr/Mr. /Ms.
(Project Coordinator)
Date……………………………
ii
DECLARATION AND COPYRIGHT
I ALLEN E DIOCLES, declare that this is my own original work and that it has not been
presented and will not be presented to any other institute/learning institution for similar or any
other Ordinary diploma award.
Signature….......……………………
This project work is a copy right protected under the Berne convention , the copy right act 1999
and other international and national enactments in that behalf, on intellectual property
It may not be reproduced by any means in full or in parts, except short extracts in fair dealings
for research or private study, critical scholarly review or discourse with an acknowledgement
without the written permission of the unit of research, consultancy and short course on behalf of
both the author and the mineral resources institute.
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ACKNOWLEDGEMENT
It has been a great experience working on a subject such as improvement of ventilation condition
in underground mine at Bulyanhulu Gold Mine Limited. I am grateful to my industrial supervisor
Eng. Abeli Kingu, Eng. Fadhili and Eng. Nelison Naforo at Bulyanhulu Gold Mine. It has been a
pleasure working in such a distinctive research with my college supervisor Eng. Khamis
Kamando who has sent me to the field work and I contributed fully to all activities especially
concern with my project in ventilation with important findings.
The feeling of a great honor has always been effective for being privileged enough to study at the
department of Mining and Mineral Processing Engineering, Mineral Resources Institute (MRI).
It is always a compliment to be thankful to all ventilation engineers to for their guidance in the
field research. I would like to give my kind thanks to ventilation office member’s engineers for
accepting to be a member of the committee, for the very important contributions to the study and
being kind enough to attend to the defense meetings coming from industry by organizing his
schedule.
Special thanks also should be given to my current and previous colleagues, for the
encouragement to perform this study during the course of the research in the interval of June to
July 2012. I would like to thank industry for giving permission to use the different equipment for
data collection given in this study.
It is a great opportunity to pay respect to my lovely act mother; Alphoncina M Rweyemamu and
my family who support me all the time being at the field work in Kahama district for great
establishment would have been achieved; in case this study could bring new insights to the
mining industry due to being one of the pioneering statistical methods in improvement of
ventilation condition in underground mine.
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DEDICATION
I would like to dedicate this research project to my young sister and brother Dellyphina Diocles
and Enock Diocles respectively, who inverses me to take this course of mining engineering, as
well as my relatives who supports me fully to continue with my study.
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ABSTRACT
This Project means to comprise Air supply in comparison with Mining Planning Methods and
other operations taking place at Bulyanhulu gold mine as well as people’s safety, when
conducting Industrial practical training purposely to be familiar and competent of underground
operations and Industrial mining activities in general.
The main activities that takes place at BGML underground are face drilling and blasting, ore and
waste development, mucking and haulage, shaft operations, mining methods such as
Conventional Cut and Fill, Almak, Long hole mining and water pumping of collected mine
drainages.
Furthermore there is a problem of the system that ventilate underground particularly zone six, as
the system needs more attention because no job will be environmentally productive and healthy
will be achieved if the air is not sufficient and no one will remain in the underground that should
vacate immediately if the system fails, as the system mainly composed of blower fans, exhaust
fans and refrigeration plant.
We need keeping on following good working procedure in suppressed dust like down enough the
muck piles before mucking, wet drilling, fixing leaking vent duct, maintenance of water sprayer
in the portal and in the ramps and control number of equipment in working areas for effective
ventilation. In order to achieve proper air supply at the mine bottom at Bulyanhulu Gold Mine
Limited (BGML) the amount that supplied over the area it should be regulated so that small
quantity of air has reported as head loss which is varies directly with the resistance at the air pass
ways, this help to get correct quantity of air that needed in a stopes with respective to the total
quantity of air that supplied early to the area concern.
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ABBREVIATION
m -meter
RL -reduced level
m3/s -meter cubic per second
FWVS2A -far west vent system number two
FWVS1A -far west vent system number one
WRAR -west return air
ERAR -east return air
CVR -central vent rise
vii
Table of Contents
CERTIFICATION ......................................................................................................................................... i
DECLARATION AND COPYRIGHT ......................................................................................................... ii
ACKNOWLEDGEMENT ........................................................................................................................... iii
DEDICATION ............................................................................................................................................. iv
ABSTRACT ................................................................................................................................................. v
CHAPTER ONE ......................................................................................................................................... 1
1.0 Introduction ........................................................................................................................................... 1
1.1 Background ....................................................................................................................................... 1
1.3 Problem statement ............................................................................................................................ 3
1.6 Objectives........................................................................................................................................... 3
1.6.1 Main objective ............................................................................................................................ 3
1.6.2 Specific objectives ...................................................................................................................... 3
CHAPTER TWO ........................................................................................................................................ 4
2.0 LITERATURE REVIEW .................................................................................................................... 4
2.1 Ventilation ......................................................................................................................................... 4
2.2 Types of ventilation ........................................................................................................................... 4
2.2.1 Primary ventilation .................................................................................................................... 4
2.2.2 Secondary ventilation ................................................................................................................ 6
2.3 Principle of mine ventilation ............................................................................................................ 6
2.3.2 Descentional ventilation ............................................................................................................. 8
2.4 Fans .................................................................................................................................................... 9
2.4.1 Main fans .................................................................................................................................... 9
2.4.2 Booster fans ................................................................................................................................ 9
2.4.3 Auxiliary fans ............................................................................................................................. 9
2.5 Ventilation appliance ........................................................................................................................ 9
2.5.1 Auxiliary fans ............................................................................................................................. 9
2.5.2 Ventilation door (Air lock) ...................................................................................................... 10
2.5.3 Pressure release flap ................................................................................................................ 10
2.5.4 Water trap ................................................................................................................................ 10
2.5.6 Regulators ................................................................................................................................. 11
2.6 Stops and Development ventilation appliance .............................................................................. 12
viii
2.7.1 Stope ventilation ....................................................................................................................... 12
2.7.2 Development end ventilation ................................................................................................... 12
2.8 VENTILATION SURVEY IN THE UNDERGROUND. ............................................................ 15
2.9 MINE VENTILATION NETWORKS .......................................................................................... 19
� Series network ............................................................................................................................. 19
� Parallel Network ......................................................................................................................... 19
2.10 REGULATION OF FANS PARAMETERS .............................................................................. 21
2.10.1 FAN DESCRIPTION ............................................................................................................. 21
2.10.2 FAN INPUT POWER CURVES .......................................................................................... 23
CHAPTER THREE .................................................................................................................................. 25
3.0 METHODOLOGY ......................................................................................................................... 25
4.0 DATA COLLECTION, ANALYSIS AND INTERPRETATION ........................................ 27
4.1 DATA COLLECTION ............................................................................................................... 27
4.2 DATA INTERPRETATION .......................................................................................................... 32
4.3 DATA ANALYSIS .......................................................................................................................... 35
5.0 RESULT DISCUSSION ........................................................................................................... 37
CHAPTER SIX ........................................................................................................................................... 38
6.0 RECOMMENDATION AND CONCLUSION ...................................................................... 38
6.1 RECOMMENDATION .............................................................................................................. 38
6.2 CONCLUSION...................................................................................................................... 38
7.0 REFERENCE ............................................................................................................................ 39
1
CHAPTER ONE
1.0 Introduction
1.1 Background
The Bulyanhulu Gold Mine is located 45km south of Lake Victoria, in the Kahama District of
the Shinyanga Region, within the Sukama tribal region of northern Tanzania (Fig 1). There are
road accesses from Mwanza, 127km to the northeast and from the town of Kahama, 84km to the south.
Fig 1: Map show location of Bulyanhulu Gold Mine
Super Ramp development is advancing upwards from the zone A-3870m RL Level as well as
downwards from the zone 1-3980m RL up to 3800m RL which will be main ventilated system.
This breakthrough will have great repercussion to the ventilation and therefore necessary to plan
well ahead changes which will have to come about so as to ensure flow changes are having
positive effect to the all areas which will be adversely affected by this breakthrough. Super Ramp
expected to be 62m3/s and deep Central Ramp was expected to downcast almost about 40 m
3/s
between 3980m RL and 3870m RL. Below A-3870m RL the ramp section to downcast only
about 20m3/s, the rest going down the alimak raises to 3800m RL. Zone A-3800m RL West
Return to exhaust about 120m3/s, this is total from Deep Central
2
At present in Bulyanhulu mine, the downcast system comprises the main shaft (60m3/s), box-cut
(304m3/s), refrigeration feeder system near to main shaft (238m
3/s) and reef 2 vent shaft
(156m3/s). Main up-cast and return air it comprise of five sub ways like far west vent
(FWVS2A, 2000KW, 280m3/s), far west vent (FWVS1A, -1.5m
3/s), west return air raise
(160m3/s), central ventilation raise (CVR) up-cast fan (77m
3/s) and east air return (ERAR,
227m3/s). Hence the total air enter in the mine is about 758m
3/s and the one leaving is 739m
3/s
that bring the difference of about 19m3/s. But for ventilation on bottom level they system of
ventilation which is called deep central ventilation which depend on main ramp toward 3800m
RL level and super ramp. Where through the ramp all access and ore drift get air at the bottom
level. Hence due to this they have to make sure that the quantity of air flow is enough to sustain
all operation that continues. After the reached on 3800m RL level it follow up-cast vent of far-
east ventilation fan (FWVS2A)
Fig no 1: Ventilation long section-primary flows for July (Bulyanhulu gold mine limited 2012)
1.2 Other researchers and approaches to tackle this issue
In order to tackle this problem we tried to use different method like collected data from field of
work, one of the project done at Bulyanhulu was done by Bluhm Burton Engineering (PTY)
Limited (BBE) which called Bulyanhulu phoenix phase 3 ventilation and refrigeration
requirements on June 2006 for Barrick gold.
FWVS2A FWVS1A WRAR CVR Box cut Main shaft ERAR
Reef 2 vent
3
1.3 Problem statement
At Bulyanhulu Gold Mine Limited the operation conducted underground which is about 1.2km
from the surface level which has pressure 88kpa and altitude is 1172 from mean sea level.
The main activities conducted are drilling, blasting, material haulage and mining services where
all activities depend on ventilation system as part of mining services.
The problem uncounted at Bulyanhulu Gold Mine Limited due to ventilation system at 3800m Rl
lead to minimization of machine life, insufficient cooling system, inadequate fume gas removal
system and little air circulation. The reason for this phenomenon might be inadequate air
produced by fans, uneven distribution of produced air might be the cause as well as more than
single operation in same area and mining planning.
1.4 Hypothesis
If we use one of air law which state that “always air takes shortest route” (Atinkson
mwaka) then it will lead to sufficient ventilation system at Bulyanhulu Gold Mine
Limited (BGML) at 3800 level.
1.5 Research question
Will law of air which states that “always air takes shortest route” lead to poor ventilation
system at Bulyanhulu Gold Mine Limited (BGML) at 3800 level? (Atkinson's, mwaka)
1.6 Objectives
1.6.1 Main objective
Improvement of underground ventilation conditions at Bulyanhulu Gold Mine Limited at 3800
reduced level.
1.6.2 Specific objectives
Review of deep central ventilation system of Bulyanhulu Gold Mine Limited (BGML)
Review of mining planning (short term and long term)
Suggestion of possible solutions to overcome problems.
4
CHAPTER TWO
2.0 LITERATURE REVIEW
2.1 Ventilation
Mine ventilation is the continuous of adequate and qualitative air to all parts of the mine
underground, where people are required to travel or work. This continuous supply of air is
required to:
� Supply oxygen for breathing purpose and must be above 19% by volume.
� Remove heat and provide comfortable working conditions and hence improve
production.
� To dilute and remove noxious and flammable gases that may be encountered during
mining operation
� To dilute and remove hazardous airborne pollutants created by various mining
operations underground example dust, fumes, aerosols, vapour etc. (The mine ventilation
society of south Africa – January 2010, Pg 22)
All this reasons above are creating and maintain an underground working environment mining is
conducive to productivity, health and safety of people. In case to archive the stated advantage the
mine fan (or fans) can create this pressure differential either by blowing air into the mine or
exhausting air from the mine. An exhaust fan pulls or sucks old air out of the exhaust airway.
This pulling causes a pressure differential which, in turn, pulls fresh air into the mine's intake.
Blower fans are used mostly in mines having little overburden. Because these mines may have
surface cracks, a blower fan is used so that any air that leaks through the cracks will leak away
from the mine, not into the mine. In many cases, one main fan is used to ventilate the entire
mine. In some large multi-level’ mines, booster fans installed on certain levels are used along
with the main fan to maintain the correct ventilation throughout the mine. If rescue teams are
working in a mine with several booster fans, they should be aware of this.
2.2 Types of ventilation
Ventilation is divided in two main types depend on the case from the place inlet air to the
working stops. In ventilation engineering we classify it as primary and secondary ventilation.
2.2.1 Primary ventilation
The basis of effective ventilation of underground mines is the adequacy of the primary
ventilation system which is the total volume flow through the mine which is conducted through
the major underground workings, normally involving splits into parallel circuits.
Factors which determine total primary volume capacity (and pressure) requirements for a mine
include the extent and depth of the mine, the complexity, and the stopping and extraction
systems, together with the size of development openings and the equipment used. One of the
major constraints on primary ventilation volume which is sometimes not adequately provided for
at the design stage is intake air capacity. Whereas high air velocities may be tolerable in return
airways and exhaust rises and shafts, (where no personnel are exposed), there is a practical limit
5
to tolerable air velocity in main intakes (shafts and declines) and main development openings
where persons travel and work. Dust generation is one problem deriving from intake velocities in
excess of 6m/sec. Moreover, high velocities require high pressure gradients and very high power
costs to maintain them.
A further major consideration with deep and extensive underground mines is the tendency to lean
towards series ventilation circuits. According to Moshab (1997) the main problem with series or
parallel-series circuits is progressive contamination of the air by recirculation from secondary
ventilation system returns, and the increased fire risk, in that the fumes and smoke from any fire
in the intake or any upstream section of the mine will be carried into working sections
downstream. In most cases, the system should be designed and scheduled to provide parallel
paths from the primary fresh air intakes through operating areas to return airways connecting to
exhaust rises and shafts. In general terms the shorter and more direct the ventilation circuit
through each working area, the better the system. Maximum use of parallel paths will reduce the
overall mine resistance for a given air flow, which in turn greatly reduces the power required and
the operating cost. The essential provision to this is that adequate volume flow through each
working area is maintained to dilute dust and contaminants and ensure operator comfort and
many mines rely on exhaust fans to provide the ventilation as it is relatively simple and easier to
regulate than a combined pressure/exhaust system.
It is strongly recommended that as part of the initial design of any mine or a planned upgrade
that computer simulation of the ventilation network be done to assist in:
� Fan selection based on fan curves.
� The effect of ventilation changes over the life of the mine. This should include start up
and completion of mining and any interim times of significance, for example at time of
maximum production.
� Selection of locations for doors, booster fans and regulators.
� Location of the second means of egress and its effect on ventilation example ladder ways
in shafts.
As the mine develops and new stopping areas are opened up, the total system alters continuously.
In any given system, primary air flows can be controlled by regulating, (closure or restriction of
some paths), or by boosting flow through designated circuits by the use of circuit fans, usually
installed on the exhaust side. Regulating flows is simpler to do and less costly, but increases the
mine resistance and reduces total primary flow. Local circuit (booster) fans increase the total
primary flow, and generally operate at high volume and low pressure, with a correspondingly
lower power demand.
6
2.2.2 Secondary ventilation
Secondary ventilation refers to the provision of ventilation to development ends, stops and
services facilities which constitute secondary circuits tapped off the primary circuit or main
through flow of air. These may be “dead end” in configuration, or be “parallel or “series in
parallel” circuits. According to Moshab (1997) the use of secondary ventilation fans and ducting
is normally required, most commonly in a "forced air" configuration, but pressure/exhaust
overlap or total exhaust may also be used. Effective secondary ventilation can be established
only if the primary ventilation system itself is adequately designed and operated. The two
systems are in fact an integrated whole hence unbalanced primary and secondary combination
can cause re-circulation, which is inefficient and potentially hazardous.
Correct selection of fans for secondary ventilation on the basis of performance characteristics
and ducting types used is critical to both the maintenance of health and safety and of efficiency
of operation. The following should be considered:
� Proper selection of fan based on duct diameter, length and type and fan duty. Fan curves
must be used to enable correct selection of the fan.
� Location of fan to prevent recirculation and damage from equipment.
� Availability of sufficient power to start and run the fan.
Some the stopping is exceed 500m and we need to overlap duct, attention to the correct design of
fan/duct combinations is essential where large volumes are required over extended distances to
cater for large scale diesel equipment. It is cost effective to provide twin ducts and two fans in
such situations, rather than to increase fan power to force larger volumes through a single duct at
the much higher pressures required. The power cost can be reduced by 50% and the reduced
pressure on the ducting greatly reduces leakage at joints and seams. The power cost saved
rapidly offsets the cost of a second fan and the additional ducting, particularly when the system
is to be split to service two or three workplaces. The application of properly engineered design to
both primary and secondary systems will enable safe and healthy conditions to be maintained,
and contamination reduced to levels which are as low as reasonably achievable. Commensurate
operational efficiencies will be maintained.
Hence according to Moshab (1997) said that the optimal layout of secondary ventilation systems
to eliminate or minimise recirculation is of fundamental importance.
2.3 Principle of mine ventilation
The fresh air ventilating a mine enters at the downcast shaft, is drawn through the working place
where it become contaminated and is removed from the mine via the up-cast shafts. A type of
mine has one or more downcast shafts where the fresh air from surface enter the mine, intake
downcast shafts through which the air flow to the workings. Fans are used to exhaust air through
the mine since natural ventilation is normally inadequate and unreliable.
7
Fig no. 2: Show the principle of mine ventilation.
2.3.1 Ascentional ventilation
Ascentional ventilation is the most common method of ventilation. Fresh air is taken through the
downcast shaft, directly to the bottom levels of the mine and then allowed to up-cast through the
working areas. The turning air is transported with in return airway on the top and out to surface
via the main up-cast fans.
Fig 3: Show the ascentional ventilation system
8
2.3.2 Descentional ventilation
This method is not recommended, especially when the presence of flammable gas is known. It is
known fact that flammable gas roof layering can occur against the ventilation flow as a result of
this gas specific gravity (lighter than air) characteristic. Thus, before down casting air thorough
workings, this phenomenon should bear in mind.
Descentional ventilation is opposite to the Ascentional ventilation because here air is taken from
down cast to shaft to the top level of the min, and then allowed to downcast through the
workings. The return air is then transported through the return airways on the bottom level to the
up-cast shaft and out surface via the main up-cast shaft fan.
9
2.4 Fans
There two main types of fain in use, radial flow (centrifugal fan) and axial fans. Generally fan
can divided in three types as follower:
2.4.1 Main fans
These are normally twins installation situated on surface at the top of the up-cast shaft, and
usually handle the bulk of the air passing through the mine that is they handle large quantities of
air. Almost all fans are centrifugal backward bladed type, with a non-overloading characteristic.
This main installation should be;
a) Regularly checked and maintained.
b) Equipped with temperature trips
c) Equipped with a pressure recording device
d) Fitted with manometer and inclined monometer
e) Fitted with a telephone
f) Able to accommodate quality measuring device
g) Vibration trips
2.4.2 Booster fans
These are installed at selected place underground to assist the main fans in handing the additional
pressure requirements as a result of increased resistance. They are sometimes up to 2 metres in
diameter and handle 70m3/s plus at 3000pascal. They should be equipped with built in
manometer and a pressure recording device.
2.4.3 Auxiliary fans
These are used to ventilate any working area not in through ventilation example development
end, dead ends, some stops, pumps chamber, dam, filter units, underground workshops and
stores and for cooling coils.
2.5 Ventilation appliance
A mine is always divided in the ventilation district and total volume of air down cast must be
distributed and controlled between these various ventilation districts or section. As air also take
shortest route or path of least resistance. Effective maintenance is required so that to reduce
resistance. As mentioned above, different ventilation appliance are utilized and installed
underground on a mine, to distribute and control the available air. This is essential to ensure an
adequate air supply to all working, where people are required to work and travel is mentioned.
2.5.1 Auxiliary fans
This type of fan is usually axial flow electric driven fans, ranging from 308mm to 760mm in
diameter and power rating from 4kw to 45kw. It used to ventilate the area that there is no natural
air flow is not automatically. In mines where drilling and blasting is done and large amounts of
10
dust are produced, auxiliary ventilation systems are often used to control and direct airflow to or
from the mining area. These auxiliary systems usually consist of small portable fan and tubing,
sometimes referred to as vent bag or fan line. Sometimes auxiliary fans are used without any
tubing to direct air into a raise. The auxiliary fan can be used to either exhaust or force the air.
The tubing, which is usually suspended from timbers or eye-bolts, carries the air to or away from
the mining area. This tubing can either be rigid (for exhausting systems) or collapsible (for
forcing systems).
Hence simply in auxiliary fan air is entering axial and leave in the same form as entrance. And it
used to ventilate in the development end, dead ends, some stops, pumps chamber, dam, filter
units, underground workshops and stores and for cooling coils.
2.5.2 Ventilation door (Air lock)
Ventilation doors are installed in series to form an airlock at various places underground. At
Bulyanhulu they applied they applies ventilation door in different level for the purpose of
controlling the amount of air that entering the workings.
The airlock requirements are:
It must not leak excessively
It must be self –closing and kept closed at all times
The installation must be as such that on only one door can be opened at any time
(interlocking)
It should be robust, strong and easy to open
Each door must be equipped with proper handle on both side pressure release flap and an
effective water trap.
It should be painted with black and yellow chevron lines or any other colour so that can
be visible.
2.5.3 Pressure release flap
A pressure release flap should be installed on all ventilation door and should be large enough to
equalize the pressure across the quickly, to facilitate the easy opening of such a door. It is
worthwhile to realize that a pressure release flap cannot fulfil its function if there is excessive
leakage through the door.
2.5.4 Water trap
All air locks through which drain pass, must be equipped with effective water trap. Water trap is
a device for allowing water to flow through an airlock without allowing air to leak through. This
device is designed on the principle of a vertical manometer. In every case there should be
different water level so that to equal the pressure across the door.
11
The following features are common to all water trap design;
� The sump must be large enough to enable mud settlement to be cleared from both sides of
the partition.
� The trap between the bottom of the partition and the sump should be large to allow an
accumulation of mud.
� The sump must be deep enough to allow trap to function when no water flow in the drain.
� The plan of the large pressure it should be large in plan than the low pressure side.
2.5.5 Stopping/walls
They installed in working to stope or block the floe air completely and can be divided as
temporary stopping, permanent stopping, and explosion- proof stopping.
1. Temporary stopping
They constructed with timber or plastic sheeting, conveyor belting, ventilation curtains
etc. This type of stopping is mainly used when temporary medication to the ventilation
system is required or for test purpose underground during air flow test, also was installed
up to moment is replaced by permanent stopping door.
2. Permanent stopping
These are concrete walls or concrete bricks, water trap (100mm diameter pipe on flow)
and gas (25mm diameter pipe against hanging wall) should fitted through all permanent
stopping’s to cater for any possible accumulation of water and/or flammable gas,
respectively behind these stopping.
3. Explosion-proof stopping
These stopping are building when section of a mine need to be sealed off to a fire or
when flammable gas (CH4) is known to be present and the possibility of a flammable gas
explosion exists. Candidates is advised to make themselves full conversant with the
standard and code of practise applicable to their mine in respect of this type of stopping,
as there are various ways of constructing explosion-proof stopping in cool and gold
mines.
2.5.6 Regulators
A regulatory is an opening in a stopping that will allow a predetermine volume of air or specific
quantity to pass through the regulator. Regulator it increase the resistance of the system in which
it is installed and hence uses up some of the available ventilating pressure, which result in the
decrease of the air volume. There is different method of regulating the air quantity that are
required to flow with in working place, some of different types of regulator are slotted rail,
sliding shutter and pipe method.
12
2.6 Stops and Development ventilation appliance
The following are appliance that used to utilize to distribute and control the air from the intake of
the working to the working face. The type of stop and development appliance at Bulyanhulu was
air-movers (venture) or ventilation duct. Air mover are sometimes used to ventilate area like
corner working place, prospects, winch chamber and places where auxiliary fans with ducting
cannot at some. The duct that used were used is 800mm by 1000mm in diameter according to the
width of the stop to ventilate.
2.7 Ventilation of working place
2.7.1 Stope ventilation
Stopes are the most important working place in gold mine as this is where the reef is mined.
More worker are employed in a stopes than in any other type of working and its essential that
adequate quantities of air are provide and controlled to maintain safe and healthy condition. The
major problem in stope ventilation much available air as possible is directed and kept on the area
of work and little air as possible should be allowed to leak into worked out area. It has been
shown that productivity performance is directly affected by environment where the bulb
temperature is 320C air velocity id 0.5m/s, the percentage performance of worker it will be 83%.
Where if the wet bulb temperature is kept constant and velocity increased to 4m/s there would
increase in performance to 95% in efficiency. Air it should prevented from flow in wrong
position by using brattices and effective strike and dip walls or curtains in such a way that, the
worker drive the maximum benefit from it.
Ruled to direct air onto the face are;
Strike walls or curtains should be kept as close as possible to the face without
affecting the air flow by any means with maximum distance of 9m from the face.
Ventilation door should be installed in correct position in did-gullies and travelling
ways.
Accumulation of rock rubble restricting the air flow should be prevented at the stopes
intake and return, as well as all faces.
Dip-gullies and face that are no longer required should seal off.
2.7.2 Development end ventilation
A development is the tunnel shaped excavation driven into virgin ground, with no natural
through ventilation and without a second outlet or escape way. They can horizontal (example
crosscut, haulage, drives etc.), inclined (example raise, box holes, shafts etc), declined (example
winzes, shafts) or vertical example shaft. Always development ventilation have no through
ventilation, they have to be ventilated by mechanical means that is with the aid of auxiliary fans
and ventilation ducts or pipes so that heat and airborne pollutant should be carefully exercised.
13
Development end ventilation has three categories as follow;
1. Forcing system
Air is required from a point at least 10 meters upstream in the nearest through ventilation
and with the aid of a fan forced through the ventilation duct and discharge to with 12
meters from the face.
- Advantages
- Air flows to the face at high velocity and good quality sir deliver at the face.
- The air is discharge to the face where worker they benefits with maximum air.
- Single fan and column are required for its installation.
- The fan and motor are in good condition hence less wear and tear on the fan.
- Leakage in the column is outward and hence easily detectable.
- Disadvantages
- Person travelling or working in the return do so in contaminated return air
from the face.
- A long re-entry period is required; hence it is unsuitable for mulit-blast
development ends.
- The return air usually flows back into the general air stream and cause
contamination.
2. Exhaust overlap system
An exhaust fan, at least 10m from in through ventilation is installed and connected to
ventilation column, which is extended into development end up to 30mfrom the face. A
second small fan is installed, 10m upstream from the exhaust column intake. A
ventilation column is fitted to this fan and extended to within 10m from the face, to force
air onto the face. The force must not handle more than two-thirds of the exhaust fan
intake volume, to ensure an adequate volume of air, and hence air is maintained in the
overlap section. It is also important that the force fan must be electrically interconnected
with the exhaust fan. This is to ensure, that should the exhaust fan fail or stop, the force
fan will then also stop, to prevent any re-circulation of the force fan.
- Advantage
- Person travelling and working in the development end so in fresh air as the
return air is inside the exhaust pipe.
- Short re-entry period are possible when used in multi-blast or high speed
development ends.
- Return air is under control.
14
- Disadvantage
- Intake air moves slowly along the drive and picks up heat dust and loco fumes
on the way. Hence the supplied to the face is interior in quality compared to
that supplied by the force system.
- Two column and two fans are required
- Poor conditions can exist in the overlap section:
1. Danger of gas accumulation here
2. Hence overlap distance in excess of 10m and air velocities above 0.3m/s
- Leakage in the exhaust column is inward, hence not easily detectable.
3. Exhaust systems
When this type is installed in through downstream from development end break way. A
ventilation column is attached to this fan and extended into the development end, up to as
close as possible to the face to exhaust air from the face. As this system does not
effectively ventilate the face, it’s not commonly used.
- Advantage
- Person working in or travelling in the end, away from the face derive
maximum benefit from the fresh air.
- Return air can be controlled
- Single fan and column is required
- Disadvantage
- Face is not effectively ventilated; therefore a gas build up at the face can
easily occur.
- Fan is situated in return air, increase chance of methane ignition and results
more wear and tear on the fan.
- Quality and quantity of the face air supplied to the face is poor
- Worker on the face derives the minimum benefit
- Leakage on the column is inward and nor easily detected.
These three system discussed above, can also be used to ventilate sinking shafts, and a fourth
system is discussed below to ventilate sinking shafts.
15
2.8 VENTILATION SURVEY IN THE UNDERGROUND.
Comprehensive ventilation surveys are necessary to determine if the mine ventilation system
meets statutory requirements, to decide what improvements in the current ventilation system are
needed, and to enable planning for future expansion. Routine measurements made to check on
the air quantity in a split or the amount of methane in the workings does not qualify as
comprising a ventilation survey. Four major areas are included under the general heading of
ventilation surveys: (Mining engineering handbook-Pg 1086)
1 Air quantity
2 Barometric Pressure
3 Air velocity
4 Temperature.
2.8.1 CATEGORY OF MINE GAS
According to Howard L Hartman 3rd
edition (1997 stated the above mine gas as shown below
- Explosive gas
- Poison Gas
2.8.1.1 EXPLOSIVE GAS
� Hydrogen(H2)
Properties
- Flammable in the range of 4.1%-74%
- Violent explosion over 7%- 8% Concentration
Sources in Mines;
- Pottassic seams
- Batteries charges
- Action of water or steam on hot materials
- Acid action on metals (iron, steel)
Effect to a human being;
Asphyxiate at high concentration
16
� Ammonia (NH3)
Properties;
- Colorless acute smell
- Pungent smell
- Smelt after blasting with ammonia explosives
- Density 0.596
Main source
- Disintegration of Nitrogen Compounds.
Effect to a human being;
- Intensive irritation of eyes
- Nose and throat produce coughing.
� Heavy hydrocarbons
Most Hydrocarbons encountered in mines are;
- Ethane (C2H6)
- Propane(C3H8)
- Butane(C4H10)
Main Source
- Mining in poorly metamorphosed coals
- Blasting works
� Acetylene (C2H4)
Properties
- Specific gravity 0.91
- Explosive Range 2.4-83%
- Ignition Temperature 3050C
Sources
- Blasting works(rarely)
- When Methane heated in low oxygen atmosphere produce acetylene.
� Methane (CH4)
Properties
- Explosive Range 5-15% with a minimum of 12.5% oxygen,
- Mixture of 0-5% not explosive but will burn near a hot source.
- Specific Gravity of 0.55 found back or roof,
- Largest component of Fire Damp 70%-80%
- Ignition temperature 6500C-750
0C
17
Main sources
- Formation of coal seams
- Metamorphism of the original organic matter
- Increasing of pressure and Temperature during coalification
-
2.8.1.2 POISONOUS GASES
� Carbon monoxide
Characteristics/Properties
� It’s both flammable and explosive.
� Ignition temperature 630 ̊C to 180 ̊C.
� S.G = 0.97
� Explosive range conc. 12.5% - 74%
Sources
- Incomplete combustion of organic based materials.
- Product of detonated explosive and diesel engines (incomplete
detonation).
- Highly toxic to body.
- CO quickly bonds with body’s hemoglobin reduce body ability to carry
oxygen
- Low temperature oxidations.
- Mine fires.
-
Effect to human being
10 – 20% Tightness across fore head, slight headache, tiredness 70% - 80%,
Respiratory failure, death.
� Oxides of nitrogen
Properties
� Non flammable
� Very Irritating
� Heavy than air
� Reddish brown
Sources
- Diesel engines
- Incomplete detonation.
Effect to human being
- At high concentration i.e. 200 – 700 ppm – fatal
18
� Sulfur dioxide (SO2)
Sources
- Blasting rocks in Sulphuric rocks
- Mine fires
- Internal combustion engines
- Some mineral springs
- Massif rocks
Effect to human being
- Very Irritating to the mucous membranes and causes muscular weakness and
fainting
- In concentration of 400 ppm to 500 ppm life threatening – dangerous to life.
� Hydrogen sulfide
Sources
- Rock massif
- Mineral sources
- Decomposing organic materials, decaying mine water which contain
sulphidic rocks
- Mine fires
- Blasting, burning of detonating cord
- Sometimes noticed near stagnant pools of water underground
Properties
� Explosive range 4.5 – 45% forms a flammable mixture in air.
� Rotten eggs smell at low concentration 0.0001%
� S.G = 1.19
� High soluble
Effect to human being
- Short-lived breathing exposure to H2S concentration of 0.1% could
cause death.
19
2.9 MINE VENTILATION NETWORKS
It is the net of connected heading through which air is flowing. Two basic circuits or
combination of airways-series or parallel – are used to distribute air through the mine but
headings could be connected in;
- Series
- Parallel
- Diagonal or complex
� Series network
A ventilation system is called series connected if air stream flows through it without splitting. In
other word in series combinations, the air ways are connected end to end and the same quantity
flows through each of the airway
i.e. Q = AV = Constant
• For series networks, flow rate in individual heading is the same Q = Q1 = Q2
• When stream flows through headings, it loose part of its head in overcoming the
resistance of individual, that is total depression is equal to the sum of individual
depression
H = H1 + H2
H = R1Q12 + R2Q2
2
Where:
Q1 - quantity of air
H1 - head/pressure
R1 – Resistance
• Total (equivalent) resistance of series connected headings equals to the sum of the
resistance of Individual headings.
R = R1 + R2
� Parallel Network
It is a ventilation system in which airflows through several branch of headings which have two
end connections. Therefore the pressure difference between the ends of each airway is the same
parallel networks are commonly employed in mine ventilation because;
1) Fresh, uncontaminated air is delivered to the workplaces on each split and;
2) The power cost is reduced sharply for a given quantity of air. It is an objective of mine
ventilation to provide a separate split of air for each workplace, where this is not practical
or possible the number of workings per split should be kept to a minimum.
Types of parallel connections;
i) Closed simple parallel connection – E.g. mine of inclined shaft
ii) Closed complex parallel connection – E.g. upper level mining in parallel field
iii) Open parallel connection
20
In general formula; for calculating
i) Air flow in parallel
Q1 = QT .
(1 + √R1/R2 + √R1/R3 + …. + √R1/Rn)
Q2 = QT .
(1 + √R2/R1 + √R2/R3 + …. + √R2/Rn)
Q3 = QT .
(1 + √R3/R1 + √R3/R2 + …. + √R3/Rn)
Then,
QT = Q1 + Q2 + Q3 + …. + Qn
ii) Total (equivalent) resistance (R) of parallel connections
1/√R = 1/√R1 + 1/√R2 + 1/√R3 + …. + 1/√Rn
iii) The mine ventilation law
The mine ventilation law is illustrated,
Mathematically as;
H = Pressure = RQ2
H = RQ2
Where: H – head/pressure
R – Resistance
Q - Quantity
21
2.10 REGULATION OF FANS PARAMETERS
The behavior of a fan under changing head-quality conditions is predictable from characteristic
curves. However there are certain variable other than the flow and resistance of the system that
exert a considerable effect on fan performance. These variables are fan rotation speed n, fan size
(diameter) D, air specific weight W, and in case of axial flow fans, the blade pitch.
The fan law:
The fan laws apply to all types of fans, regardless of location with respect to the system (blower,
exhaust or booster). They are summarized in the following table.
Variance in
performance
characteristics
Law 1, with speed change, n
(D and W constant)
Law 2, with size
change, D (W and
n, D Constant)
Law 3 with
specific changes
W (n and D
Constant)
Quantity Directly As square Constant
Head, H As square Constant Directly
Power, Pa or Pm As cube As square Directly
Efficiency n Constant Constant Constant
Where n = fan rotation
D = fan size
W = air specific weight
2.10.1 FAN DESCRIPTION
A fan is an appliance that converts mechanical energy delivered to the fan, into potential energy
(pressure) and kinetic energy (velocity).
Pressure of cause is necessary to overcome the resistance of a particular duct or system in which
the fan is operating
Fans may vary in diameter, power rating, air volume handled and pressure created.
Fan can be divided into main fans, Auxiliary fans and booster fans.
Main types of Fans
Main Fans
These fans are normally situated on surface at the top of the up – cast shaft
that exhaust air through the mine, and is commonly known as the “lungs” of a
mine. These fans can either be electrical driven axial flow or centrifugal fans
that can handle volumes of air ranging from 25-450 m3/s at pressure of 400 Pa
to 9,0kpa and power ratings from 50 KW to 4100 KW.
22
1. Auxiliary Fans
These fans are used underground to ventilate working places not in through
ventilation (i.e where air would not natural enter) such as developments ends,
back stopes, up – cast, workshops etc. Usually, these fans are electric driven
axial flow fans, that can handle air volumes of up to 15m3/s @2,1 KPa, power
ratings, from 4KW to 45 KW and varies in diameter form 380 mm to 760 mm,
but other sizes may exist.
2. Booster Fans
The installation of booster underground sometimes becomes unavoidable.
Longer airways have to be serviced when workings approach the extremities,
resulting in an increased resistance of a mine/shaft system. In order to
overcome the increase resistance of such system, booster fans need to be
installed to assist the main fans. Booster fans generally handle between 60 and
90m3/s at pressures ranging up to 30KPa. These installations can either be
axial flow or centrifugal fans but axial flow or as normally little or no
additional excavation are required for installation purposes.
One important factor to be considered when installing these units, are that
they should be installed as close as possible to the up – cast shaft to prevent
recirculation’s, which can contaminate intake air with heat, blasting fumes and
noxious gases from fires. (The mine ventilation society of South Africa
January- 2010).
23
2.10.2 FAN INPUT POWER CURVES
• Over loading power curve
Is a power curve where the power increases continuously as the quantity increases
until eventually the power become the high and the fan will trip or burn out given the
sketch below illustrates Cleary about over loading fan.
• Non – overloading Power curve.
Is a power curve where the power increases as the quantity increases up to a certain
point when the power will slowly decrease while the quantity continue to increase
The sketch below illustrates.
24
25
CHAPTER THREE
3.0 METHODOLOGY
Both quantitative and qualitative techniques were used to collect data of temperature profile
(Ambient) Air velocity profile, pressure profile (Barometric pressure) and gas profile to obtain
actual ventilation networking of mine particularly at the deep of the mine. Standard and close
observations of the primary long section system, air flow, fan locations systems were also
performed interviewing of the competent operators in the working places.
Secondary ventilation survey were measured area of working places, air flow velocity,
temperature (Dry and wet bulb), relative humidity, Barometric pressure and gases.
Instruments used to measure ventilation mine parameters were;
� Multi-gases monitor – this was used to measure gases in the underground areas.
� Personal emergency device – for carbon monoxide gas monitor
� Whirling hygrometer – for measuring wet bulb temperature dry bulb temperatures
� Anemometer vane probe – to measure air Velocity
� Digital hygrometer – temperature measurements.
� Notebook and forms provided – to record data’s obtained
� Densitometer – for measuring stope heights and widths.
Procedures used to measure ambient temperatures and velocity.
1. Ambient temperatures
� Poor some water in the tube – left hand side
� Whirling the instrument for 30 seconds
� White end of thermometer side used to read wet bulb temperature and red end part of
thermometer used to read dry bulb temperature
� Record in the note book.
2. Air velocity
� Keeping Anemometer vane probe upward against air flow while on the
footwall drift access or stope access.
� Record the data obtained in the notebook
26
THE PRIMARY FLOWS – VENTILATION LONG SECTION
This is one of the methods I used to study qualitatively and quantitatively the underground
ventilation systems at Bulyanhulu gold mine and come up with couple of findings which helped
me to perform this project. It shows the fresh air intake and return air which is contaminated after
being used.
27
CHAPTER FOUR
4.0 DATA COLLECTION, ANALYSIS AND INTERPRETATION
4.1 DATA COLLECTION
The data was obtained during the field that conducted to accomplish the research project, but
before we tried to conduct to obtain data we as the team advances through the mine during
exploration, all ventilation controls should be checked, especially those in the affected part of the
mine. When you come to a regulator or door, the position of it should be noted on the map by the
map man and it should be reported to the command centre. The command centre should be told
the type of damage you find and the extent of the damage. For example, if a bulkhead or other
type of structure has been blown out by explosive forces, you should note the direction in which
it appears to have blown.
Data collected was based on three parametric measure of ventilation system that found
Bulyanhulu mine such as temperature (heat stress), silica dust content in air and DPM test result.
4.1.1 Temperature (heat stress) result
Heat stress can be defined as environment measurement of air temperature; air flow, the level of
heat exchange and metabolic rate of person so that to maintain constant body temperature of
370C due to having great regulating mechanism.
According to P. du Toit (2007) ventilation objectives guidelines are;
1.2.1 Stopes
� Wet bulb temperature between ….27.5-29.50C, not exceed 32
0C
� Air velocity ………………………0.25m/s (minimum),
� Dust……………………………….below 1mg/m3.
1.2.2 Development
� Wet bulb temperature ……………27.5-29.50C, not exceed 32
0C
� Air quality deliver ………………..0.15m3/s/m
2 (minimum),
� Dust ………………………………below 1mg/m3
According to the guidelines above was stated to be worked at performance of 100%, but for
Bulyanhulu mine its wet bulb temperature was 280C to 31.5
0C with draw and dust (silica
content) was 0.05.mg/m3.
28
For the success of this project following are the findings and information obtained when
performing my project from different levels and location in the underground
1 EAST RETURN AIR RAISE
Fan velocity pressure 0.2kpa
Fan static pressure 2.1kpa
FAN TOTAL PRESSURE 2.3kpa
Temperature 18.20C wet bulb/28.1
0C
Barometric Pressure 88.4Pa
Relative Humidity 43.6%
Table 1: East return air raise
2 CENTRAL VENT RAISE
Fan velocity pressure 0.2KPa
Fan static pressure 1.4KPa
FAN TOTAL PRESSURE 1.6KPa
Temperature 17.50C wet bulb/27.2
0C
Barometric Pressure 88.3Pa
Relative Humidity 41.1%
Table 2: Central vent raise
29
3 FAR WEST VENT STATION
Fan velocity Pressure -1.1KPa
Fan static Pressure 1.8KPa
FAN TOTAL PRESSURE 0.7KPa
Temperature 17.40C wet bulb/27
0C dry bulb
Barometric Pressure 88.4Pa
Relative Humidity 42.6%
Table 3: Far west vent station
4 WEST RETURN AIR RAISE
Fan velocity Pressure 2.1KPa
Fan Static Pressure 0.2KPa
FAN TOTAL PRESSURE 2.3KPa
Temperature 17.50C wet bulb/28
0C dry bulb
Barometric Pressure 88.4Pa
Relative Humidity 37.6%
Table 4: West return air raise
30
Also to achieve correct conclusion the following data were obtained from the field area at
Bulyanhulu mine at 3800 level concern with heat stress as shown below;
Area/location
Wet
bulb
(OC)
dry
bulb
(OC)
Drift dimension Velocity (m/s)
Quantity
(m3/s)
Height
(m)
Width
(m)
Area
(m2)
V1 V2 (V1+V2)/ 2
(m/s)
A-3800 HZD 31.6 36.5 5.25 6.0 31.5 0.9 0.9 0.9 28.35
A-3800 FWDE 32 34 7.1 6.1 43.31 1.1 1 1.05 45.47
A-3800 FWDW 31.5 34 7.1 6.2 44.02 0.9 1.2 0.54 23.77
A-3800 FWD
Vent access west
29
32
7.1
6.2
44.02
0.9
0.9
0.9
39.62
A-3800 Decline 28.6 32.7 7.2 6.3 45.36 1.3 1.2 1.25 56.7
A-3800-208 32 35 4.7 4.9 23.03 0.8 0.9 0.85 19.57
A-3800 O/D W 31.4 33.2 4.96 3.3 16.36 1 0.8 0.9 14.73
Table 5: show the temperature result from 16th
to 21st July 2012
Area/location
Wet
bulb
(OC)
dry
bulb
(OC)
Drift dimension Velocity (m/s)
Quantity
(m3/s) Heigh
t (m)
Width
(m)
Area
(m2)
V1 V2 (V1+V2)/2
(m/s)
A-3800 HZD 28 31.5 5.25 6.00 31.5 0.9 0.9 0.9 28.35
A-3800 FWDE 32 34 7.1 6.1 43.31 1.2 1.3 1.25 54.14
A-3800 FWDW 33 35.5 7.1 6.2 44.02 0.9 1.1 1 44.02
A-3800 FWD
Vent access west
30
33
7.1
6.2
44.02
1.1
0.9
1
44.02
A-3800 Decline 28 31 7.2 6.3 45.36 1.2 1.3 1.25 56.7
A-3800-208 34 35 4.7 4.9 23.03 1.4 0.8 1.1 25.33
A-3800 O/D W 32.5 35 4.96 3.3 16.36 1 0.8 0.9 14.73
Table 6: show the temperature result from 23rd
to 27th
July 2012
31
4.1.2 Diesel particulate matter (DPM) test result
Diesel Particulate matter (DPM) test that was conducted at Bulyanhulu mine in order to obtain
the amount of carbon in the air that produced by the machine operation which called as noxious
gases or other type of product due to diesel example soot. The test is under OH office through
providing instrument to the all machine operator and record automatically.
Crew
Machine name
Work location
TWA TC
(mg/m3)
Barrick OEL
(TC) mg/m3
Muck & haulage HT 63 (Truck) 3800E 0.2 0.160
Muck & haulage L 707 (LHD) 3800 – 190E 0.154 0.160
Waste development L 710 (LHD) 3800E 0.316 0.160
Waste development HT 56 (Truck) 3800W 0.153 0.160
Ore development L 135 (LHD) 3800-190E 0.235 0.160
Upper east L 711 (LHD) 3800E HZD 0.186 0.160
Ore development L 304 (LHD) 3800-221W 0.251 0.160
Upper east HT 58 (Truck) 3800W 0.428 0.160
Waste development HT 64 (Truck) 3800E HZD 0.222 0.160
Waste development L 711 (LHD) 3800E FWD 0.268 0.160
Table 7: Show the DPM test result from 16th
to 21th
July 2012
Where: TWA- Time Weighted Average calculated in 8 hours of exposure
TC - Total carbon
Note; Total carbon is equal to Organic carbon (OC) + Element carbon (EC)
4.1.3 Silica dust sampling result
It consist all element that was due to the dust during operation example drilling, charging, truck
operation during excavation and loading equipment. For Bulyanhulu mine they deal with silica
content because other was not produced in extent that it harmful, hence only silica dust was
produced in large and tend to increase time to time in different level due to operation conducted.
Crew
Job type/section
Work location
TWA silica (quartz)
dust (mg/m3)
OEL silica (quartz)
dust (mg/m3)
Waste development Charge up 3800W 0.02 0.05
Waste development Charge up 3800E 0.03 0.05
Waste development Jumbo 3800-221 0.01 0.05
Waste development Supporter mine 3800-231E & W 0.02 0.05
Waste development LHD operator 3800-241 0.04 0.05
Waste development Truck operator 3800-231 0.02 0.05
Table 8: Shows the silica dust concentration at 3800m reduced level.
32
4.2 DATA INTERPRETATION
Below is the graph that shows the temperature variation in different stopes at 3800m reduced
level, with the Occupational Exposure Limit for 8 hours shift that are recorded from 16th
up to
21st July 2012 and from 23
rd up to 27
th July 2012.
Graph 1 show the exposure temperature in different stopes at 3800m reduces level from 16
th to
21st July 2012 versus dry bulb Occupational Exposure Limit
Graph 2 shows the relation of exposure temperature in different stopes from 23
rd to 27
th July
2012 versus dry bulb Occupational Exposure Limit
A-3800 HZDA-3800
FWDE
A-3800
FWDW
A-3800
FWD Vent
access west
A-3800
DeclineA-3800-208
A-3800 O/D
W
Dry Bulb Temp 36.5 34 34 32 32.7 35 33.2
Dry Bulb OEL 32 32 32 32 32 32 32
29
30
31
32
33
34
35
36
37
Tem
pe
ratu
re (
oC
)
EXPOSURE TEMPARATURE IN DIFFERENT STOPES
A-3800
HZD
A-3800
FWDE
A-3800
FWDW
A-3800
FWD
A-3800
Decline
A-3800-
208
A-3800
O/D W
Dry Bulb Temp 31.5 34 35.5 33 31 35 35
Dry Bulb OEL 32 32 32 32 32 32 32
28
29
30
31
32
33
34
35
36
Tem
pa
ratu
re (
0C
)
EXPOSURE TEMPERATURE IN DIFFERENT STOPES
33
Also the amount of Diesel Particulate matter (DPM) in mine is taken in consideration because
the amount of DPM produced it varies with the amount of heat absorbed and generated to the
surrounding. Each machine have certain fixed ratio of its temperature for fuel burning hence if
any phenomenon occur it alter the output after the combustion that leads to such Diesel
Particulate Matter to be generated. Below the graph shows the DPM produced in different stopes
and its variation as what is Occupational Exposure Limit that measured in Time Weighted
Average calculated in 8 hours of exposure.
Graph 3 shows the average diesel particulate matter at 3800m reduced level versus Barrick
Occupational Exposure Limit
Muck
&
haulag
e
(3800E
)
Muck
&
haulag
e
(3800 -
190E)
Waste
develo
pment
(3800E
)
Waste
develo
pment
(3800
W)
Ore
develo
pment
(3800-
190E)
Upper
east
(3800E
HZD)
Ore
develo
pment
(3800-
221W)
Upper
east
(3800
W)
Waste
develo
pment
(3800E
HZD)
Waste
develo
pment
(3800E
HZD)
Total Carbon (mg/m3) 0.2 0.154 0.316 0.153 0.235 0.186 0.251 0.428 0.222 0.268
Barrick OEL (mg/m3) 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16 0.16
0
0.05
0.1
0.15
0.2
0.25
0.3
0.35
0.4
0.45
Die
sel p
art
icu
late
ma
tte
r (m
g/m
3)
AVERAGE DIESEL PARTICULATE IN DIFFERENT STOPES AROUND
3800m REDUCED LEVEL
34
Personal dust sampling was done on July 2013 from different working groups underground in
waste development areas to determine silica (quartz) exposure. Sample were collected and sent
to SKC South Africa for silica analysis using NIOSH 7602 Method. The average results were as
shown in the graph below:
Graph 4 shows the average silica dust exposure versus Barrick Occupational Exposure Limit
Charge up
(3800w)
Charge up
(3800E)
Jumbo
(3800 -
221)
Supporter
mine (3800
- 231E &
W)
LHD
operator
(3800 -
241)
Truck
operator
(3800 -
231)
Silica dust exposure (mg/m3) 0.02 0.03 0.01 0.02 0.04 0.02
Barrick OEL (mg/m3) 0.05 0.05 0.05 0.05 0.05 0.05
0
0.01
0.02
0.03
0.04
0.05
0.06E
xp
osu
re le
ve
l (m
g/m
3)
AVERAGE SILICA DUST EXPOSURE
35
4.3 DATA ANALYSIS
From the graph above the data obtained are analysed according to the percentage as follow:
� From graph number 1,
Average Barrick Occupational Exposure Limit in temperature = 320C
Average dry bulb temperature:
(36.5 + 34 + 34 + 32 + 32.7 + 35 + 33.2)0C
7
(237.4)0C
7
= 33.90C
• Percentage increase in temperature from the Barrick Occupational Limit is given by:
33.90C X 100%
320C
1.059 X 100%
105.9% - 100%
Hence the increase in temperature from the Barrick OEL is 5.9%
� From graph number 2,
Average Barrick Occupational Exposure Limit in temperature = 320C
Average dry bulb temperature:
(31.5 + 34 + 35.5 + 33 + 31 + 35 + 35)0C
7
2350C
7
= 33.570C
• Percentage increase in temperature from the Barrick Occupational Limit is given by:
33.570C X 100%
320C
1.049 X 100%
104.9%– 100%
Hence the increase in temperature from the Barrick OEL is 4.9%
36
� From graph number 3
Average Barrick Occupational Exposure Limit for diesel particulate matter = 0.16 mg/m3
Average diesel particulate matter (DPM):
0.2 + 0.154 + 0.316 + 0.153 + 0.24 + 0.186 + 0.251 + 0.428 + 0.222 + 0.268
10
Average DPM = 0.2418mg/m3
• Percentage increase of diesel particulate matter from Barrick Occupational Exposure
Limit:
0.2418 mg/m3 X 100%
0.16 mg/m3
1.5112 X 100%
151.12% - 100%
51.12%
Hence the increase in diesel particulate matter from the Barrick OEL is 51.12%
� From graph number 4
Average Barrick Occupational Exposure Limit for silica dust exposure = 0.05 mg/m3
Average silica dust exposure:
0.02 + 0.03 + 0.05 + 0.05 + 0.05 + 0.05
6
Average DPM = 0.042 mg/m3
• Percentage increase of diesel particulate matter from Barrick Occupational Exposure
Limit:
0.042 mg/m3 X 100%
0.05 mg/m3
0.84 X 100%
84% - 100%
= -16%
Hence the decrease in silica dust exposure from the Barrick OEL is -16%
37
CHAPTER FIVE
5.0 RESULT DISCUSSION
According to graph number 1, the maximum temperature was 36.50C at 3800m reduced level in
horizontal drive (HZD) and minimum temperature was 320C at zone A- 3800m reduced level in
forward drive (FWD) vent access, and its average exposure temperature was 34.250C west.
Whole average of the data reading of dry bulb temperature was 33.90C, which exceeds the
normal exposure temperature at Bulyanhulu by 5.9%.
According to graph number 2, the maximum temperature was 35.50C at 3800m reduced in
forward drive west (FWDW) and minimum temperature was 31.50C at 3800m reduced in
horizontal drive (HZD) and its average exposure temperature was 33.50C. Whole average
temperature was 33.50C, which it exceeds normal planned of 100% temperature by 4.9%.
According to graph number 3, maximum diesel particulate matter (DPM) was 0.428 mg/m3
at
3800m reduced level west in of carbon content and minimum was 0.153 mg/m3 at 3800m
reduced level east. Its average exposure was 0.29 mg/m3 while the total average of Diesel
Particulate Matter (DPM) exposure calculated was 0.24 mg/m3 which exceed by 51.12% from
the normal diesel particulate matter limit set by Bulyanhulu mine.
According to graph number 4, silica dust content it contain crystalline silica which later form
fibrosis (scar tissue) in the lungs which reduce the ability of the lungs to extract oxygen from the
air we breathe. Maximum silica result was 0.04 mg/m3 and minimum was 0.01 mg/m
3 and its
average exposure was 0.02 mg/m3. While the total average of silica dusts exposure was
0.04mg/m3 which are less from the silica exposure limit by 16%.
38
CHAPTER SIX
6.0 RECOMMENDATION AND CONCLUSION
6.1 RECOMMENDATION
Silica dust content it contains crystalline silica which later form fibrosis (scar tissue) in the lungs
which reduce the ability of the lungs to extract oxygen from the air we breathe. At Bulyanhulu
mine the data was collected it show on how 38000m reduced level it have maximum temperature
average of about 34.50C, maximum diesel particulate matter (DPM) of about 0.29 mg/m
3 but
minimum silica content exposure of about 0.02 mg/m3 in mine. We need keeping on following
good working procedure in suppressed dust like down enough the muck piles before mucking,
wet drilling, fixing leaking vent duct, maintenance of water sprayer in the portal and in the ramps
and control number of equipment in working areas for effective ventilation.
6.2 CONCLUSION
In order to achieve proper air supply at the mine bottom at Bulyanhulu Gold Mine Limited
(BGML) the amount that supplied over the area it should be regulated so that small quantity of
air has reported as head loss which is varies directly with the resistance at the air pass ways, this
help to get correct quantity of air that needed in a stopes with respective to the total quantity of
air that supplied early to the area concern.
In mining planning either long term or short term they should plan the mine can consume the air
quantity which it can consumed by the all machine, people and other mechanical means on the
mine without any scarcity on the area which later at same mining zone if more than one mining
was conducted and different machine running they consume much air quantity that at the end
leads to heating of air which due to high air resistance air tend to circulate at the same area or
zone.
Avoid installing the auxiliary fan at the ramps which have no sufficient air supply of Bulyanhulu
occupational exposure limit and hence still formulate more heat generation on the mine stopes,
which also caused by long ventilation duct during air transmission to the new stopes. Any major
change to the ventilation system should be modeled prior to the change being implemented. This
is so the modeling will confirm the effect of the change on all ventilation splits in the site and
that all relevant standards can be maintained.
39
7.0 REFERENCE
� Howard L. Hartman (1992) SME Mining Engineering hand book 2nd
edition vol. I&II
SME Inc.
� H.L.Hartman (1990) SME Mining Engineering hand book, New York, Society of Mining
Engineers American Institute of Mining Metallurgical and Petroleum Engineers.
� Mutmansky, J.M. and Wang, W.H. (1997) "Results of Field Studies on Stratification of
Diesel Particulate Matter in Mine Openings," Proc. 6th Int'l Mine Vent Cong., Ramani,
R.V.,ed., SME, Littleton, CO, pp. 155-162.
� Ramani, R.V. (1992) "Chap. 11.6: Mine Ventilation," SME Mining Engineering
Handbook, 2nd
Ed., Vol. 1, Hartman, H.L., et al., ed., pp. 1052-1092.
