2. LETTER OF SUBMISSION
The University Of Zambia,
Kwacha hostels, Block 5 room 1,
P.O Box 32379, Lusaka.
July, 2014.
The Head of Department,
Mining Engineering Department,
P.O Box 32379, Lusaka.
Dear Sir,
REF: SUBMISSION OF PROJECT REPORT AS PARTIAL FULFILMENT FOR THE
AWARD OF THE DEGREE OF BACHELOR OF MINERAL SCIENCE IN MINING
ENGINEERING (B.Min.Sc).
With reference to the above, I, Tembo Misheck, Computer Number 29047749 and National
Registration Number 178190/54/1 do hereby submit this report titled “Propose ideal ventilation
system for Mufulira Mine Deeps Section from 1340 to 2020 metre levels”.
The main objectives of this project are; (1) to determine the total mine heat load for the Deeps
section based on planned production, machinery and workforce from 1340 to 2020 metre levels,
(2) to determine air quantity required to dilute the total mine heat load, (3) to propose the main
intake and return airway systems for the entire Deeps section and (4) to recommend appropriate
type of fans to be used in the Deeps section.
The information of the survey was obtained through underground fieldwork as well as verbal
discussions with the mine ventilation officers, Mufulira Mine and literature was researched
through, Journals, Internet, textbooks and lecture’s notes. I hope and trust that this report will
meet the requirements of the department.
Yours faithfully,
Tembo Misheck.
CONTACTS: Cell: +260976944015, +260966944015
Email: misheckctembo@yahoo.com
4. ii
ACKNOWLEDGEMENT
I would like to express my gratitude to all members of staff in the Mining Engineering
Department for their valuable support and guidance during my period of study. My sincere
thanks go to my supervisor Dr. V. Mutambo. Without his guidance the production of this
report would have been difficult.
My sincere appreciation also go to the management of Mopani Copper Mines (Plc.) for having
given me an opportunity to do this project. I am greatly indebted to the following people for the
assistance they rendered to me during my attachment at Mopani Copper Mines (Plc.): Mr. V.
Deon (Superintendent mining training), Mr. J. Lesa (Head of mining training), Mr. A. Mwansa
(senior mine technical officer), Mr. H. Mubita (senior mine technical officer), Mr. Ken C.
Simfukwe (ventilation engineer), Mr. A. Matanga (senior geologist), Mr. J. Kamukwamba
(senior ventilation officer), Mr. C. Sinyinza (senior ventilation officer), Mr. T. Kamanga (senior
ventilation officer), Mr. J. Ndalama (senior ventilation officer), Mr. E. Mulenga (senior
ventilation officer) and my project mentor Mr. D. Mpongwe (senior ventilation officer).
Immense and heart felt appreciation go to my family for supporting me financially, spiritually
and emotionally.
I would also like to thank my classmates for making my stay on campus worthwhile and
memorable.
To every story there are people we forget to mention but strictly speaking, they may be the
backbone of who we turn out to be. Through this passage I may not list all of them but what I
cannot forget is to thank them all, thank you all!
5. iii
ABSTRACT
Mufulira mine has embarked on the expansion program by extending the life of the mine. This
involves the mining of the lower echelon from 1340 to 2020 metre levels here in referred to as
‘Mufulira Mine Deeps Section’. However, since the commencement of mining, Deeps section
has been experiencing high temperatures in the working places caused by mine heat build-up.
The working conditions are further worsened by leakages of ventilation ducts, increased air
resistance in airways and recirculation of air. These high temperatures reduces productivity and
has an adverse effect on the mining cost, health and safety of underground personnel.
Therefore, this thesis is aimed at optimizing the current ventilation system and determine the
ventilation requirements for mining below 1340m level down to 2020m level. Specifically, the
study established air quantity required to dilute the total mine heat load which is 1619 m3
/s.
Furthermore the study has proposed appropriate type of fans, ventilation circuit, refrigeration
plant, main intake and return airways.
6. iv
TABLE OF CONTENTS
DECLARATION……….……………………………………………….……………………………………..i
ACKNOWLEDGEMENT……………………………………………………………………………………..ii
ABSRACT……………………….……………………….……………………………………………………iii
CHAPTER ONE: INTRODUCTION
1.1 Location of the Study Area………………………….……………………………………………………1
1.2 History…………………………….……………………………………………………………………... 1
1.3 Problem Statement……………………………………………….……………………………………….2
1.4 Objectives……………………………………………………………………………………………….. 3
CHAPTER TWO: MINE VENTILATION LAYOUT
2.1 Main Intake Air Network………………………………………………………………………….……...4
2.2 Main Return Air Network…………………………………………………….…………………………..5
2.3 Primary Access………………………….……………………………………………………………......6
2.4 Current Mining Method…………………………………………………………………………………..7
CHAPTER THREE: LITERATURE REVIEW
3.1 Mine ventilation………………………………………………….……………………………………… 8
3.2 Adequate ventilation………………………….…………………………………………………………. 8
3.3 Air distribution and control…………………………….……………………………………………….. 9
3.4 Recirculation and leakage of air……………………….……………………………………………….. 10
3.5 Sources of mine heat……………………………………………………………………………………...10
3.5.1 Auto Compression……………………………...……………….………………………………….……10
3.5.2 Rock strata………………………………………..….………….………………………………….……11
3.5.3 Diesel powered equipment…………………...…………………………………………………….…… 12
3.5.4 Electrical equipment…….………………………..….………….………………………………….……13
3.5.5 Human bodies…………………….………….....……………….………………………………….……14
3.5.6 Oxidation of minerals and timber….…………...……………….………………………………….……14
3.5.7 Lamps……………………………………………..….…………………………………………….…… 14
3.5.8 Hot pipes and electrical cables.....…………...…………………………………………………….……. 15
3.5.9 Blasting operations…….………………………..….………….………………………………….…….. 15
7. v
3.5.10 Movement of rock strata………………………..….…………………………………………….……..15
3.5.11 Mine fires.....…………...…………………………………………………….…………………………15
3.5.12 Hot water fissures…….………………………..….………….………………………………….……..16
3.6 Summary of heat sources……………………………………….………………………………………. 16
3.7 Heat management controls………………………………………………………………………………. 17
CHAPTER FOUR: METHODOLOGY
4.1 Desktop Study…..…………………………….…………………………………………………………..18
4.2 Measurement of air velocity in airways….……………………………………………………………….18
4.2.1 Vane Anemometer.....…………...…………………….……………………………………….…………18
4.2.2 Dust Dispenser.....…………...………………………………………………………………….………..20
4.3 Area measurement of an airway…………………………….…………………………………………… 20
4.4 Determination of air quantity………………………………………….………………………………….21
4.5 Measurement of temperatures…………………………………………………………………………… 23
4.5.1 Whirling Hygrometer…………...…………………….……………………………………….…………23
4.6 Measurement of virgin rock temperature………………………………………………………………... 24
CHAPTER FIVE : DATA ANALYSIS
5.1 Heat Load from Auto Compression……………………………………….……………………………...27
5.2 Heat Load from Rock Strata…………………………….………………………………………………..28
5.3 Heat Load from Diesel Powered Equipment………………………………….………………………….31
5.4 Heat Load from Electrical Equipment…………………………………………………………………... 33
5.5 Heat Load from Humans………………………………………………………………………………….. 36
5.6 Total Mine Heat Load………………………….…………………………………………………………37
5.7 Air quantity required to dilute the total mine heat load…………………………………………………. 38
5.8 Calculation for refrigeration ventilation requirement…………………….………………………………38
CHAPTER SIX: VENTILATION REQUIREMENTS
6.1 Proposed ventilation circuit…..………………….………………………………………………………. 39
6.2 Proposed main intake airways between 1340m and 2020m levels……………….……………………... 40
6.3 Proposed Main return airways between 1340m and 2020m levels…......................…………………….. 40
6.4 Proposed refrigeration plant...………………………….………………………………………………... 41
6.5 Proposed underground fans……………………….…………………………………………………….. 42
8. vi
CHAPTER SEVEN: CONCLUSION AND RECOMMENDATIONS
6.1 Conclusion……………………………………….………………………………………………………. 43
6.2 Recommendations………………………………….……………………………………………………..43
REFFERENCES...………………………………………………….………………………………………….45
APPENDIX I...………………………………………………….…………………………………………….. 46
APPENDIX II...………………………………………………….…………………………………………….47
APPENDIX III………………………………………………….……………………………………………...48
LIST OF TABLES
Table 2.1: Existing main shafts and Intake Airway capacities……………………….....……………………..4
Table 2.2: Quantity of air handled by Upcast Shafts……………………………………….………………….6
Table 3.5.5: Metabolic heat………………………………….………………………………………………... 14
Table 3.6: Heat source and relative contribution to the total heat load of mine…………………….…………16
Table 3.7: Heat management controls………………………………………………………………………... 17
Table 4.4: Air velocities, air quantity and cross section areas of various airways…………………….………22
Table 4.5: Wet and dry bulb temperatures…………………….……………………………………………….24
Table 4.6a: VRT measurements at 1440m level……………………………………………………………….25
Table 4.6b: VRT measurements at 1340m level….…………………………………………………………... 26
Table 5.2a: Virgin rock temperature (VRT)……………………….…………………………………………..29
Table 5.2b: VRTs and their corresponding levels……………………………………………………………. 29
Table 5.2c: Heat generated from rock strata at different levels……………………….……………………….31
Table 5.3: Heat generated by diesel equipment…………………….………………………………………….32
Table 5.4a: Fan characteristics………………………………………………………………………………... 33
Table 5.4b: Overall total fan power developed……………………………………………………………….. 33
Table 5.4c: Conveyor heat load……………………….……………………………………………………….34
Table 5.4d: Mass conveyed…………………………………………………………………………………… 34
Table 5.4e: Conveyor availability…………………….………………………………………………………..35
Table 5.5a: Metabolic heat……………………………………………………………………………………. 36
Table 5.5b: Manpower at deeps section………………………………………………………………………. 36
9. vii
Table 5.6: Total mine heat load………………………………………………………………….……………. 37
Table 5.7: Psychrometric properties………………………………………….………………………………..38
Table 6.2: Main intake airways……………………………………………………………………………….. 40
Table 6.3: Main return airway raises……………………………………………………………….…………. 41
LIST OF FIGURES
Figure 1.1: Location of Mufulira Mine……………………………………….……………………………......1
Figure 2.1: Intake ventilation circuit………………………………………………………………………...... 5
Figure 2.2: Return ventilation circuit…………………………………………………………………….…….6
Figure 2.3: Decline layout in the deeps section…………………………………………………….………….7
Figure 2.4: Mechanized Continuous Retreat………………………………………………………………...... 7
Figure 4.3: Across section area of an arched tunnel…………………………………………………………...21
Figure 6.1: Proposed ventilation circuit…………………………………..………………................................39
Figure 6.4: Figure 6.4: Surface plant and bulk air cooler……………………………………………………...41
Figure 6.5: Booster fan performance curve…………………………………………………………………....42
LIST OF GRAPHS
Graph 5.2a: Geothermal temperature gradient……………………………………………................................29
Graph 5.2b: Heat production curve for the deeps section………………………………………………………30
10. 1
CHAPTER ONE: INTRODUCTION
1.1 Location of the Study Area
This project research was conducted at Mufulira mine. Mufulira mine is located on the
Copperbelt Province of Zambia in the town of Mufulira. Mufulira license is mined in three
geographical areas namely Mufulira West Portal, Mufulira East Portal and the main Mufulira
mine which comprises of the upper, central and deeps sections.
Figure 1.1: Location of Mufulira Mine - Google Earth, Map of Zambia
1.2 History
The mineral deposit was discovered in 1923 along the Mufulira River by James Moir and Guy
Bell, two prospectors employed by Rhodesian Congo Border Concession limited. By 1930
drilling had indicated ore reserves in excess of 100,000,000 tonnes at an average grade of 4.4%
11. 2
copper. By 1931 a small concentrating plant had been completed. However, before production
could begin, the operations in the mine were suspended due to industrial depression which took
place during World War I. However, the mine was re-opened in 1933 and by October that same
year the underground workings were dewatered and metallurgical plant construction started. In
January 1937, the smelter was completed and production of blister copper started. By 1938, the
mine’s annual production of blister copper had reached 60,000 tonnes.
Prior to political independence two mining firms operated in Northern Rhodesia. These were the
Rhodesian Select Trust and Anglo American Corporation (AAC). In 1972 the Zambian
government acquired equity holding and the two units were renamed Roan Select Trust (RST)
and Nchanga Consolidated Copper Mines (NCCM) and operated under Zambia Industrial and
Mining Corporation (ZIMCO). In 1982 RST and NCCM were merged to form the giant Zambia
Consolidated Copper mines (ZCCM), in which the Government of the Republic of Zambia
(GRZ) held 60.3% share and AAC through Zambia Copper Investment (ZCI) held the balance.
In the year 2000, the Zambian government privatized the mines. Two major mining companies
emerged after the privatization namely Glencore Ltd owned Mopani Copper Mines (MCM),
Nkana and Mufulira assets. Then AAC owned Konkola Copper mines (KCM).
1.3 Problem statement
Mufulira mine has embarked on the expansion program being called “Mufulira Mine Deeps
project”. This project is intended to cover from 1340m level to the proposed deepest 2020m
level. The Deeps project was commenced in the mid part of the year 2004 with the expected ore
annual production of two million tonnes by the year 2015.
However, since the commencement, Mufulira Mine Deeps section has been experiencing high
temperatures in the working places caused by mine heat build-up. The working conditions are
further worsened by leakages of ventilation ducts, increased air resistance in airways and
recirculation of air. These high temperatures reduces productivity and has adverse effects on the
health and safety of underground personnel. Therefore, this study is aimed at evaluating and
proposing ideal ventilation system for Deeps section in order to ensure adequate ventilation in
compliance with chapter nine of 1974 mining regulations of the laws of Zambia.
12. 3
1.4 Objectives
The main objectives of this project are:
To determine the total mine heat load for the Deeps section based on planned
production, machinery and workforce from 1340 to 2020 metre levels.
To determine air quantity required to dilute the total mine heat load.
To propose the main intake and return airway systems for the entire Deeps section.
To recommend appropriate type of fans to be used in the Deeps section.
13. 4
CHAPTER TWO: MINE VENTILATION LAYOUT
2.1 Main Intake Air Network
Intake air is supplied into the mine through six sub vertical shafts. Three shafts are located on the
western side of the mine (i.e. numbers 11, 12, and 14 shafts) extending from surface to 810m
level. The other three shafts are located on the eastern side of the mine (i.e. numbers 5, 7 and 9
shafts) go down to 500m level. While west and east portals are additional intake airways
extending from surface to 500m level. The air to the lower levels is supplied via:
Matelo sub inclined shafts Matelo one (M1), Matelo two (M2) goes down to 1040m
level and Matelo three (M3) to 880m level.
Musombo SV shaft from 500m level to 1400m level.
14 X 3 conveyor tunnel from the base of number 14 shaft to 1040m level.
Plan ‘C’ conveyor and service tunnel from the base of M1, M2 and 14 X 3 tunnels at
1040m level down to 1400m level.
56P8 Fresh Air Intake raise from 880m level to 1423m level.
56 Decline Ramp from 1340m level to 1457m level.
A new 6.7m diameter hoisting shaft which is being sunk will have a capacity to carry 318 m³/s at
12 m/s. The shaft will extend from surface to 2020m level.
Table 2.1: Existing main shafts and Intake Airway capacities
Shaft 5# 7# 9# 11# 12# 14# Total
Diameter or
(W x B)
5.4 x 1.9 6.0 x 4.2 6.7 6.7 6.7 5.5
Area m² 10.2 25.2 35.2 35.2 35.2 23.7
Velocity m/s 8 8 10 10 10 10
Nominal
Volume m³/s
82 202 352 352 352 237 1577
14. 5
Figure 2.1: Intake ventilation circuit – Mufulira Mine Planning Department
2.2 Main Return Air Network
There are two main upcast shafts:
Number 8 shaft; 6.7m diameter. This is a series of stepped shaft extending from surface
to 865m below surface.
Number 10 shaft; 6.7m diameter. This shaft extends from surface to about 350m below
surface where it splits into two separate shafts which extend to 865m.
Existing surface upcast fans
At number 8 shaft.
Three surface fans in parallel rated 212 m³/s at 5 kPa
Variable speed motors 575/497 rpm
Motor rating are 1350/820 kW
At number 10 shaft
Two surface fans in parallel rated 287 m³/s at 2.75 kPa
Single speed motors 595 rpm
Motor rating 1350 kW
15. 6
Table 2.2: Quantity of air handled by Upcast Shafts
Figure 2.2: Return ventilation circuit – Mufulira Mine Planning Department
2.3 Primary Access to Deeps Section
Primary access to the Deeps section of Mufulira mine is a single 5m wide x 6m high decline. The
decline serves as the main fresh air intake as well as the main men, material and rock
conveyance. The decline is inclined at 8 degrees and it lies 78 meters away from the orebody in a
South–West and North-East orientation and perpendicular to the principal stress axis. In this
position only the south portions of the decline will be exposed to principal stress. The decline has
been designed in such a way as to maximize truck haulage cycle times.
Shaft Number Air Quantity
8 636 m3
/s
10 575 m3
/s
6 100 m3
/s
16. 7
Figure 2.3: Decline layout in the deeps section – Mufulira Mine Planning Department
2.4 Current Mining Method
Currently Mufulira mine Deeps section is conducting mining using mechanized continuous
retreat (MCR) method. This mining method is essentially a sub-level open stopping method. The
mining methods involve establishing longitudinal stope blocks across the strike of the ore body.
A slot is opened to join two or three level to produce a free face from which stoping began.
Stoping began from the east and west wards’ leaving no rib pillars until the echelon is reached on
both sides.
Figure 2.4: Mechanized Continuous Retreat – Mufulira Mining Department
17. 8
CHAPTER THREE: LITERATURE REVIEW
3.1 Mine ventilation
Mine Ventilation is essentially the act of supplying air, controlling its amount and movement in
an underground mine, this air being in both good quality and quantity. The quality and quantity
of air supplied is largely depend on the efficiency of the ventilation system being employed at a
particular mine. The task of supplying good quality air in sufficient quantities is an important
element in the operation of a mine as concerned. Therefore, the mining firm must ensure that
monitoring of ventilation conditions underground is undertaken regularly not only for purpose of
compliance with the statutory requirements but also to be proactive in contributing to the
efficient running of the mine. A good ventilation system takes care of excessive heat,
temperatures, dust and gases in a mine.
The initial design of a ventilation system is a balance of many factors with health being of
paramount importance. Dust levels, temperature and air quantity surveys should be conducted
periodically to determine the conditions of the system. Planning for future exigencies is
necessary to be able to incorporate changes in the initial design. The resistance to airflow of
these systems requires the efficient maximum utilization of air volumes necessary to maintain a
safe and healthy underground atmosphere. This can be achieved only by the proper distribution
and control of adequate air volumes. Although there is no ideal or standard system of mine
ventilation, the effectiveness and efficiency of the system will be determined by how well certain
fundamentals are applied and maintained.
3.2 Adequate ventilation.
The volume of air required to adequately ventilate working areas underground depends on
factors such as the amount of explosives used, number of miners underground, quantity of heat
emitted into the ventilating air by various sources of heat and various gases emitted by the rock.
It is a requirement by the Zambia Mine Regulation to constantly supply adequate fresh air in all
working areas and traveling routes.
18. 9
Mining Regulations 902(2) considers ventilation to be adequate if it:
a) Ensures that the amount of oxygen in general body of the air is not less than nineteen per
centum by volume;
b) Ensures that the amount of carbon dioxide, carbon monoxide, nitrous fumes, sulphur
dioxide and hydrogen sulphide in the general body of the air do not exceed the quantities
set out against each such gas;
c) Dilutes or removes any other toxic gas or fume so that the amount of such gas or fume in
the general body of the air conforms to the requirements prescribed, from time to time, by
the chief Inspector;
d) Dilutes or removes any harmful dust so that the amount of such dust in the general body
of the air conforms to the requirements prescribed, from time to time, by the chief
inspector;
e) Maintains working conditions free from dangerous temperature at high relative
humidities in the general body of the air; and
f) Provides any diesel unit with not less than 0.05 cubic metres of air per kilowatt for the
purpose of diluting or removing any toxic gas or fume in the general body of the air at
places where such diesel units operates.
3.3 Air distribution and control
Air distribution refers to the supply of air in the desired amounts in different working areas in an
underground mine. It is achieved successfully by adopting a ventilation method and plan suitable
for the mining method to be employed in exploiting the mineral deposit. Effective distribution of
air ensures that both direction and quantity of air flowing are controlled. Ventilation is rendered
useless if the fresh air is not properly distributed to working places where it is required to
maintain wet-bulb temperatures below the Thresh Hold Limit value, remove dust and dilute
gases present in the area.
Effective and efficient air distribution can be achieved through optimal selection of the location
of control devices and of fans. A proper air distribution system would supply air to a specific
working place in required quantities. Controlled splitting of air is an aid in the distribution of
mine air and is aided by using control devices such as stoppings and regulators. Control devices
19. 10
in mine ventilation are used for separating the intake and return airstreams and to regulate the
flow of air in different airways.
The mine airflow distribution is completely defined by the following;
The parameters of the airways, shape, area, length and characteristics of the airway surface
The layout of the mine openings
Sources of pressure in the system , for example fans
The intersection between the airways, mine openings and pressure sources
3.4 Recirculation and leakage of air
Recirculation of air is usually caused by leakage of return air into fresh air, and results in an
excessive load on the fan. The load on the fan can be reduced if precautions are taken to ensure
an air-tight installation. It also occurs when air is kept within a closed circuit and should not be
confused with the situation when air is reused, as in series ventilation circuits. Recirculation
provides difficulties in the control of temperature-humidity of air in hot humid mines.
Leakage is the most common cause of inefficient distribution of air in underground mines. The
most likely places for leakage to occur are at fan installations, in caved ground, in shafts between
downcast and upcast compartments, at stoppings or doors in cross cuts between adjoining
airways. In reality, ventilation leakage rates in underground operations have typical values of
30% of the total airflow, and in some cases, this is as high as 50%.
3.5 Sources of mine heat
The process of controlling high temperatures in working areas underground requires the
knowledge of the sources of heat. It is necessary to estimate the amount of heat from these
sources. Heat in deep mine will result mainly from;
3.5.1 Auto compression
The primary cause of high temperatures in deep mines is auto compression. This is the increase
of air temperature due to change in potential energy. As air travels down the intake airways from
the surface, its elevation decreases and there is a corresponding conversion of potential energy
20. 11
into enthalpy. The magnitude of the change in enthalpy can be estimated using the steady flow
energy equation for a higher elevation (Z₂) to a lower one (Z₁), assuming no heat flow and no
work done.
dH = H₂ - H₁ = g (Z₂ - Z₁)/1000 kJ/kg
Where,
dH = change in enthalpy
H = enthalpy (J/kg)
Z = elevation (m)
g = acceleration due to gravity (9.81 m/s²)
The enthalpy thus increases by 0.981kJ/kg for every 100m decrease in elevation. This means for
every 100 metres of increase in depth, auto compression adds 0.981 kilojoules to each kilogram
of air. The value of 0.981kJ/kg of air per 100 metres is constant for every mine. For dry air, the
thermal capacity is 1.005kJ/˚C and hence change in dry bulb temperature is 0.976o
c.
The change in dry bulb temperature is much less if water evaporates in the shaft.
The following equation is used in determining heat due to auto compression;
H = dt x Cp x Mₐ
Where,
H = heat due to auto compression (kJ/s)
dt = change in dry bulb temperature (°C)
Cp = thermal capacity (kJ/kg)
Mₐ = mass flow rate of air (kg/s)
3.5.2 Rock strata.
Another major source of heat in mines is geothermal energy from the rock strata. This is the most
difficult heat source to analyse and predict due to the number of variables that influence strata
heat flows. These are virgin rock temperature (VRT), thermal conductivity and diffusivity of the
rock, the age and the size of the opening, the quality and psychometric condition of airflow and
wetness factors, roughness and texture of the exposed rock.
The equation used to estimate the heat flow from the rock strata is given by:
21. 12
Q = DFA (VRT - WB) X
(perimeter)¹˙³
12
x
(KρC)ᵒ˙⁵
13 𝑋 10⁶
Where,
Q = Heat pick up at development ends (kW)
DFA = Daily Face Advance (m/day)
VRT = Virgin Rock Temperature (o
C)
WB = Wet Bulb Temperature (o
C)
K = Thermal conductivity of the rock (w/m°C)
𝛒 = Density of the rock (2646.5 Kg/m³)
C = Specific heat of the rock (KJ/Kgo
C)
The constant 13 x 10⁶ is the value of the Kpc for quartzite.
The heat production curve is plotted using the above equation and can be used to evaluate heat
generated from stopes and development ends.
3.5.3 Diesel powered equipment
The Zambia mine and mineral acts require that a mine should provide 0.05m³/s of ventilation air
per kilowatt of engine power. However, typical ventilation requirements for acceptable operation
of diesel equipment are 0.035m³/s ‘over the engine’ per kilowatt power.
Two methods are available for quantifying the heat from diesel equipment. The first one is based
on the total rated power of a fleet and the amount by which it is used.
𝑞 𝑑𝑖𝑒𝑠𝑒𝑙 =
𝑅𝑃
ɳ,
× 𝑃𝑡 𝑎𝑣 × 𝑃𝑡 𝑢𝑡
Where,
𝑞 𝑑𝑖𝑒𝑠𝑒𝑙 = Diesel heat load, kW
𝑃𝑡 𝑎𝑣 = Percent time available, %
𝑃𝑡 𝑢𝑡 = Percent time utilization, %
ɳ, = efficiency, average 33%
𝑅𝑃 = Rated Power, kW
22. 13
The second method is based on energy consumed by the fleet in the time it is used. Theoretically,
this method will produce the same answer as the first, but due to difficulty in assigning average
availability and utilization figures to the entire fleet, usually serves only to demonstrate
consistency.
3.5.4 Electrical equipment
The heat from electrical equipment will mainly be from the conveyor belts and auxiliary fans.
Auxiliary fans
The power consumed by auxiliary fans is calculated from consumed electrical power or
assuming an efficiency and using measured air quantity and pressure.
𝑞 𝑓𝑎𝑛 =
𝑄 𝑓× 𝑃 𝑓
ɳ,
Where,
𝑞 𝑓𝑎𝑛 = Total fan power, kW
𝑃𝑓 = Fan pressure, kPa
𝑄 𝑓 = Fan quantity, m3
/s
ɳ, = efficiency, %
Conveyor belts
Conveyors belts do useful thermodynamic work when increasing the potential energy of coal or
rock being conveyed. The heat load is given by the difference between electrical energy
consumed and real work done,
𝑞 𝑐𝑜𝑛𝑣 = 𝐸𝑙𝑒𝑐𝑡𝑟𝑖𝑐𝑎𝑙 𝑝𝑜𝑤𝑒𝑟 𝑐𝑜𝑛𝑠𝑢𝑚𝑒𝑑 − 𝑊𝑜𝑟𝑘 𝑑𝑜𝑛𝑒 (𝑘𝑤)
=
(𝐸 𝑙𝑜𝑎𝑑× 𝑡 𝑙𝑜𝑎𝑑+𝐸 𝑛𝑖𝑙× 𝑡 𝑛𝑖𝑙)
(𝑡 𝑙𝑜𝑎𝑑+ 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓)
−
𝑀 𝐶× 𝑔 × (𝑧2−𝑧1)
(𝑡 𝑙𝑜𝑎𝑑+ 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓) ×3.6 × 106
Where;
𝑞 𝑐𝑜𝑛𝑣 = Conveyor heat load, (kW)
𝐸𝑙𝑜𝑎𝑑 = Electrical power at average belt load, (kW)
23. 14
𝐸 𝑛𝑖𝑙 = Electrical power at nil loads, (kW)
𝑡 𝑜𝑓𝑓 = Time the conveyor is off, (hours)
𝑡𝑙𝑜𝑎𝑑= Time the conveyor runs loaded, (hours)
𝑡 𝑛𝑖𝑙 = Time the conveyor runs nil loads, (hours)
𝑀 𝐶 = Mass moved in total time(𝑡𝑙𝑜𝑎𝑑 + 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓), Kg
3.5.5 Human bodies
The amount of heat produced by the human body varies depending on the amount of work
performed. According to the research carried by C. H. Wyndham in 1971, Human Sciences
Laboratory, Chamber of Mines Research Organization, the following data in Table 3.5.5 was
provided;
Table 3.5.5: Metabolic heat
Amount of work Heat produced (W)
At rest 90
Light rate of work 200
Moderate rate of work 275
Hard rate of work 470
3.5.6 Oxidation of minerals and timber
Oxidation of minerals (such as pyrites, sulphites) and timber may lead to the production of heat.
This heat is in turn added to the total heat load of the mine. This source accounts for a very small
percentage and often insignificant.
3.5.7 Lamps
Carbide lamps produce about 500-750kJ per hour but are no longer in common use in mines.
Electric cap lamps produce about 10kJ/hr and 100 watt electric lights about 160 kJ/hr. Heat
24. 15
generated by electric lights depend on the wattage of the lamp. Most of the energy supplied to
light bulb appears as heat. Fluorescent tubes are now in common use as they are more efficient.
3.5.8 Hot pipes and electric cables
Mufulira Mine handles many electric cables; common ones include the 11KV cables. However,
energy dissipated in the form of heat in electric cables is usually small but if the cable is too
small for the load it is carrying, the cable will tend to overheat and produce heat which is added
to total mine heat load. Compressed air pipes and return water column are usually situated in the
down cast shaft for the ease of access at all times. At the same time these pipes will often convey
air or water which is hotter than the down cast airs and some of this heat will be conducted
through the walls of the pipes into the air.
3.5.9 Blasting operations
When a high explosive is detonated energy is released and heat accounts for over 70% of energy
released. Much of this heat can be removed by ventilating air. Some of the heat enters the rock
and is dissipated there. However this is dependent on the temperature gradient existing between
the rock and the explosion heat. This may affect the cooling of the mine at a later stage. As
mentioned earlier this source of heat is not very critical as heat from it is removed by ventilating
the area.
3.5.10 Movement of rock
Gravitational forces acting above excavations produced by mining operations cause subsidence.
This action may result into crushing, fracturing and grinding within the rock mass. This
movement performs work against friction and all the work appears as heat which is kept inside
the rock. The quantity of heat generated, however is small enough to be ignored completely in
relation to the heat transferred from the rock.
3.5.11 Mine fires
Large fires produce great quantities of heat. If the volume of air feeding the fire is known and the
reduction in oxygen content due to oxygen being burnt in the fire is measured, the heat produced
can be calculated. For a fire to occur a fuel, heat and oxygen should be present. However, fires
25. 16
are most dominant in coal mines due to the presence of the mentioned factors. Oxygen is from
ventilation air, coal is the fuel itself and heat could be from any of the sources highlighted above
as applied to underground mines; this is not the case for Mufulira mine.
3.5.12 Hot water fissures
Hot water fissure has a small contribution to the total heat load of the mine, but if present in an
area it`s effect can be significant. In Mufulira mine, at 1390m level dewatering drive has hot
water coming out of the rock. This water had temperatures as high as 38o
C when temperatures
readings were taken at the time this project was conducted. This water, by virtue of coming from
the rock, had temperatures equivalent to that of the rock or exceeding it. This water transfers it`s
heat to the air as it comes out during evaporation. This in turn, increases the latent heat of the air.
3.6 Summary of heat sources
The list below summarizes the main sources of heat and gives an approximate estimate of
contribution from each source in a typical deep level mine.
Table 3.6: Heat source and relative contribution to the total heat load of mine
Source Typical relative contribution to total
heat picked up in mine
Auto compression of the air 25%
Heat flow from the rock 45%
Waste heat from machinery 15%
Heat from human bodies 5%
Other sources such like oxidation of timber and
minerals, lamps, hot pipes, electric cables, explosives
occasional source, fire, hot water fissure, ground
movement.
10%
26. 17
3.7 Heat management controls
Heat management controls are aimed at either reducing face temperatures or increasing air
cooling power to make marginally high temperatures acceptable. Two common methods are
summarised as follows;
Table 3.7: Heat management controls
Method comments
Increase face
ventilation rate
Increase face velocity for improved air cooling power. Localised
improvement can be obtained by
Increase auxiliary fan capacity and duct size
Using exhaust overlap ventilation system
Use face ventures or jet fans
Note that increasing fan power will also increase heating of air delivered
to the face.
Refrigeration
Cool all or part of the mines ventilation by refrigeration so that acceptable
temperatures are maintained underground throughout the year.
This is normally considered a last resort due to the capital and operational
costs involved. However, in deep mines (1000 + m) the effect of heat
from the rock strata and heat due to auto compression may make
refrigeration inevitable.
The practical limit for ventilating a deep, hot mine before resorting to refrigeration is one cfm
per tonne of ore mined per year. (1 cfm =4.74 X 10-4
m³/s).
27. 18
CHAPTER FOUR: METHODOLOGY
4.1 Desktop study
In order to fulfill the objectives of this project desktop study was used and the comprehensive
literature review involved the usage of:
Journals
The Internet
Textbooks
Interviews
Field works:
This was done by physically going underground and checking the existing ventilation conditions.
Detailed survey:
Measurement of cross section areas, air quantities, velocities, pressures, virgin rock
temperatures, wet and dry temperatures were conducted in different locations in the driveways.
4.2 Measurement of air velocity in airways
Various instruments can be employed in measuring mine air velocity. During this study, a vane
anemometer in conjunction with a stopwatch was used to measure air velocities of greater than
1m/s. For air with velocities less than 1m/s, a dust dispenser was used.
4.2.1 Vane anemometer
Vane anemometer is used for the routine measurement of air velocities greater than 1m/s. This
instrument is fitted with a clutch mechanism for disconnecting the dials from the spindle. In
addition to the clutch, the instrument is provided with a zero setting device. To prevent
inaccuracies and inconsistent in readings which are as a result of the operator obstructing the
flow of air, the instrument is provided with the anemometer extension rod with a swivel head.
Measuring procedure
A measuring point is selected at a location where the airway is of regular in cross section,
reasonably straight and unobstructed. If bends or obstructions are present, measurements must be
made upstream of them.
28. 19
A chosen position is marked off at right angles to the general air direction, this mark will not be
vertical in an inclined airway but will be at right angles to the dip of the air flow. In this study, a
vane anemometer which measures air velocities in the range 1 to 15m/s was used.
With the instrument still in the carrying case, a rod is attached to the anemometer in order to
minimize the possibility of damaging the instrument. The vanes of the instrument are blown to
check if they are turning freely. Also the clutch and zeroing devices are checked before the
commencement of measuring. The pointers on the instrument are set to zero. By standing slightly
down stream of the marked position and facing a side-wall, the anemometer is positioned in the
lower corner of the measuring station with the vanes facing the air stream at right angles. The
instrument is held in the air current for a few seconds to the air flow before it is started in order
for the instrument to attain the corresponding velocity of air. In this position, the anemometer
and the stopwatch are started simultaneously and the watch is immediately released allowing it to
hang by the string.
The anemometer is traversed across the airway at a speed of approximately 0.2m/s, taking care
that it is always perpendicular to the direction of airflow. At half way across the airway, the
operator turns facing the opposite sidewall making sure that the vanes of the anemometer still
face the air stream. The traverse is completed and terminated in the other bottom corner. In this
investigation, the traversing period was taken to be 50 seconds and hence, after this period the
instrument and the stopwatch were stopped. The instrument was stopped by pressing the clutch
so that further revolutions of vanes are not recorded.
The uncorrected velocities are calculated by dividing the anemometer readings by the traverse
times, 50 seconds in this case. If the difference between two readings is greater than 5 % then
there is an error in the observations or unsteady flow conditions. When this occurs further
readings, are taken.
29. 20
4.2.2 Dust Dispenser
A dust dispenser measures air velocities of less than one metre per second. It works on a
principle of timing the speed of movement of a visible cloud of dust over a known distance.
Measuring procedure
A portion of an airway of regular cross-section and with minimum bends and obstructions is
selected, and two points 3m apart, are marked on the side-wall. An assistant provided with dust
dispenser aligns the beam of his lamp at right angles to the direction of airflow at the upstream
mark and then releases a cloud of dust in the centre of the airway at arm’s length upstream of this
mark. A stop -watch is started as the dust cloud enters the observer’s beam from his lamp which
is aligned at right angles to the downstream mark.
4.3 Area measurement of an airway
The airways at Mufulira Mine have different shapes implying that their areas are computed
according to the shape. For airways like ramps whose sections are rectangular, the area was
obtained from the product of the width and height. The two dimensions were measured using a
tape measure. The area for an arched tunnel was computed by employing the formula below;
A= (W × H) + πr2
/2
Where;
A = area of the arched tunnel
W = width of the tunnel
H = height of the tunnel
r = radius of the arch
30. 21
Figure 4.3: Across section area of an arched tunnel
4.4 Determination of air quantity
The volume of air flowing through an airway or duct at any particular point is the product of the
corrected air velocity and the mean cross sectional area at that point. The quantity of air flowing
through an air way is calculated from the following equation:
Q = V × A
Where;
Q = quantity of air in m³/s
V = velocity of air in m/s
A = airway cross sectional area in m²
The air velocities, air quantity and cross section areas of various airways measured are recorded
in the table 4.4 shown below.
31. 22
Table 4.4: Air velocities, air quantity and cross section areas of various airways
Location
Cross-Section
Area (m)
Air Velocity
(m/s)
Air Quantity
(m3
/s)
Acceptable
Velocity
Ranges (m/s)
Intake
56 Ramp 30 4.13 124 2.5 – 8.0
1357 ml
Main Haulage 24 0.79 19 2.5 – 8.0
1373 ml
Main Haulage 24 0.75 18 2.5 – 8.0
1390 ml
Return raise 4 7.5 30 10.0 – 15.0
1407 ml
Return raise 2.5 8.8 22 10.0 – 15.0
1420 ml Intake
56P8 7.1 6.48 46 4.0 – 8.0
1423 ml
Crosscut 20.25 1.25 25 4.0 – 8.0
1440 ml
Stope end 20.25 0.49 10 0.5 – 4.0
1457 ml
Develop. end 24 1.16 28 0.5 – 4.0
32. 23
4.5 Measurement of Temperatures
Measurement of the wet-bulb and dry-bulb temperatures in the air were determined by using the
whirling hygrometer.
4.5.1 Whirling Hygrometer
Whirling hygrometer consists of two thermometers, one with its wet-bulb exposed to the air and
the other with the wet-bulb wrapped up in a piece of Muslim cloth, which dips into a reservoir of
distilled water. The one exposed to air is referred to as dry-bulb and the other one wrapped into
the Muslim cloth as wet-bulb. The dry-bulb thermometer measures the ambient air temperature
and the wet-bulb thermometer wet-bulb temperature. Evaporation of the water from the wet
Muslim cloth reduces the temperature of the wet-bulb thermometer in direct proportion to the
dryness of the air, and the readings of the two thermometers gives all the information required to
obtain the relative humidity of the air from a set of hygrometrical tables.
Measuring Procedure
The whirling hygrometer is first exposed to the surrounding environment so that it acquires the
temperature of the ambient air. The instrument is held at arm’s length whilst facing the air
stream. It is then whirled at a rate of approximately three revolutions per second (3rev/s) for at
least 30 seconds to give an air velocity of about 3m/s. In areas where the velocity of the air is
more than 3m/s whirling of the instrument was not necessary, but the instrument is held normal
to the flow of the air current. This is usually done when measuring the velocity of air discharged
from the ventilation duct.
The temperature readings where read as quickly as possible. The wet-bulb temperatures were
read first in order to avoid inaccurate readings as it tends to rise when whirling ceases. Contact
with the bulb was avoided by holding the instrument by its handle when taking readings to
prevent the transmission of heat from the operator to the bulb. The instrument was kept away
from any near-by local source of heat such as from the lamps. Table 4.5.1 below shows
temperature measurements taken at different locations in the deeps section.
33. 24
Table 4.5.1: Wet and dry bulb temperatures
4.6 Measurement of virgin rock temperature
Virgin rock temperature (VRT) refers to the temperature of the Insitu rock which has not been
affected by heating or cooling from any artificial source. Virgin rock temperatures are employed
in determining the geothermal gradient of the mine. In this research, a Laser beam thermometer
was used to measure the temperature of the Insitu rock.
Location Activity
Temperature
( °C )
TLV
Wet-Bulb
(°C)Wet-bulb Dry-bulb
1340 ml
Loader Workshop Maintenance 28 32 31
1357 ml
Entrance
Waste hauling
29 33 31
1373 ml
Decline Toro lashing 29 33.5 31
1390 ml
Block 56 P8 Stope drilling 29 34 31
1407 ml
Block 57 P9
Vent wall
Construction 30 34 31
1407 ml
C58 West stope
Drilling with
Rig 59 30 34 31
1423 ml
Block 56 P8
CAT 22
Lashing 31 33 31
1423ml,Block 56 P8
Ventilation raise
Exhaust fan
Installation 32 35 31
1440ml, Block 57 P5
Footwall drive
Stope drilling
31 34 31
1440ml,Panel 56
Decline
Waste
hauling 30.5 32 31
1457ml, Panel 56
Mining drive west A Supporting 34 36.5 31
1457ml, Panel 56
Decline
Decline
drilling 32 35 31
34. 25
Measuring Procedure
A Laser beam thermometer is pointed into a borehole and then push the scan button on the
thermometer. The temperature is noted and the scan button is released. The procedure is repeated
at least five times and then calculate the average temperature value. The average temperature
value is then taken as the Virgin Rock Temperature of that location.
Distance in to bore hole against temperature
1440m level Block 56 Panel 5 Angle: +10ᵒ N/W
Wet bulb temperature 36o
C Relative humidity 100%
Borehole depth +20m Dry bulb temperature 38o
C
Table 4.6a: VRT measurements at level 1440m level
Depth from collar into hole (m) Temperature ( o
C)
1.8 38.5
3.6 38.7
5.4 39.2
7.2 39.8
9 40
10.8 40
12.6 40.2
14.4 40.4
The average VRT = 39.6°C
35. 26
Distance into borehole against temperature
1340m level Block 52/53 (Boundary) Angle: Horizontal
Borehole depth +120m Relative humidity 90%
Wet bulb temperature 30.5o
C Dry bulb temperature 34o
C
Table 4.6b: VRT measurements at 1340m level
Depth from collar into hole (m) Temperature (o
C)
1.8 34.5
3.6 35
5.4 35.5
7.2 36.3
9 37
10.8 37.5
12.6 38.2
14.4 38.5
16.2 38.8
18 39
19.8 39.2
21.6 39.5
23.4 39.6
25.2 39.7
27 39.8
28.8 40
30.6 40
The average VRT = 38.1°C
36. 27
CHAPTER FIVE: DATA ANALYSIS
5.1 Heat Load from Auto Compression
The primary cause of high temperatures in deep mines is auto compression. This is the increase
in temperature of air due to change in potential energy as air travels down an intake airway from
the surface. As air travels down a shaft from the surface, its elevation decreases and there is a
corresponding conversion of potential energy into enthalpy. The magnitude of the change in
enthalpy can be estimated using the steady flow energy equation for a higher elevation (Z₂) to a
lower one (Z₁), assuming no heat flow and no work done.
dH = H₂ - H₁ = g (Z₂ - Z₁)/1000 kJ/kg
Where,
dH = change in enthalpy
H = enthalpy (J/kg)
Z = elevation (m
g = acceleration due to gravity (9.81 m/s²)
The enthalpy thus increases by 0.981kJ/kg for every 100m decrease in elevation. This means that
for every 100 metres of increase in depth, auto compression adds 0.981 kilojoules to each
kilogram of air. The figure of 0.981kJ/kg of air per 100 metres is constant for every mine. For
dry air, the thermal capacity is 1.005kJ/˚C therefore,
Change in dry temperature =
change in enthalpy
thermal capacity
=
𝑑𝐻
Ϲ𝚙
=
0.981kJ/kg
1.005kJ/kgᵒϹ
= 0.976ᵒϹ
37. 28
The change in dry bulb temperature is much less if water evaporates in the shaft. At Mufulira
mine the change in temperature due to auto compression are 0.5°C wet bulb and 0.9°C dry bulb
temperatures per 100 metres of increase in depth.
The following equation is used in determining auto compression;
H = dt x C𝚙 x Mₐ
Where,
H = heat due to auto compression (kJ/s)
dt = change in dry bulb temperature (°C)
C𝚙 = thermal capacity (kJ/kg)
Mₐ = mass flow rate of air (kg/s)
On 1440m level H = 0.9(100/100) x 1.005 x 578 = 523 kW
On 1540m level H = 0.9(200/100) x 1.005 x 693 = 1254 kW
On 1640m level H = 0.9(300/100) x 1.005 x 798 = 2165 kW
On 1740m level H = 0.9(400/100) x 1.005 x 1274 = 4609 kW
On 1840m level H = 0.9(500/100) x 1.005 x 1429 = 6463 kW
On 1940m level H = 0.9(600/100) x 1.005 x 1584 = 8596 kW
On 2020m level H = 0.9(700/100) x 1.005 x 1769 = 11200 kW
Therefore, total auto compression heat load = 34810kW.
Mₐ = Q/ASV
ASV =
287.045(𝑡 𝑑𝑏+273.15)
P −e
𝑚3
/𝑘𝑔 𝑎𝑖𝑟
e = 610.5 × exp[
17.27×𝑡 𝑤𝑏
𝑡 𝑤𝑏+237.3
] − 0.000644 × 𝑃 × (𝑡 𝑑𝑏 − 𝑡 𝑤𝑏), Pa
e = Actual vapor pressure
5.2 Heat Load from Rock Strata.
Another major source of heat in mines is geothermal energy from the rock strata.
38. 29
Table 5.2a: Virgin rock temperature (VRT)
1340m level Block 52/53 1440m level Block 56 Panel 5
Borehole depth 120m
Angle: Horizontal
Borehole depth 20m
Angle: 10° N/W
Average VRT = 38.1°C Average VRT = 39.6°C
Geothermal temperature gradient (𝐺) =
Change in temperature
Change in depth
=
∆𝜃
∆𝑋
=
39.6−38.1
1440−1340
ᵒc/m
= 0.015 ᵒϹ/m
This translates to 1.5o
C/100m
Table 5.2b: VRTs and their corresponding levels
Level 1340m 1440m 1540m 1640m 1740m 1840m 1940m 2020m
VRT(ᵒC) 38.1 39.6 41.1 42.6 44.1 45.6 47.1 48.3
Graph 5.2a: Geothermal temperature gradient
37
38
39
40
41
42
43
44
45
46
47
48
49
1340 1440 1540 1640 1740 1840 1940 2040
Virginrocktemperature,(˚c)
Depth below surface, (m)
39. 30
The equation below was used to estimate the heat flow from the rock strata:
Q = DFA (VRT - WB) X
(perimeter)¹˙³
12
x
(KρC)ᵒ˙⁵
13 𝑋 10⁶
Where,
Q = Heat pick up at development ends in kilowatts
DFA = Daily Face Advance (m/day)
VRT = Virgin Rock Temperature (o
C)
WB = Wet Bulb Temperature (o
C)
K = Thermal conductivity of the rock (w/m°C)
𝛒 = Density of the rock (2646.5 Kg/m³)
C = Specific heat of the rock (KJ/Kgo
C)
The constant 13 x 10⁶ is the value of the Kpc for quartzite.
The heat production curve was plotted using the above equation and has been used to evaluate
heat generated from stopes and development.
Graph 5.2b: Heat production curve for the deeps section
0
0.01
0.02
0.03
0.04
0.05
0.06
0.07
0.08
0.09
0.1
0.11
0.12
30 35 40 45 50
HEATPRODUCTIONPERTONNEOFROCKBROKENPER
MONTH,(KW)
VRT, (°C)
40. 31
Assuming an average production of 167, 000 tonnes per month from development and stopes,
using the heat production curve, the heat generated was calculated as follows:-
Table 5.2c: Heat generated from rock strata at different levels
Level Production Rate
(tonnes/month)
Heat Produced Per ton/month
(kW)
Total Heat Generated
(kW)
1440mL 167,000 0.062 10, 354
1540mL 167,000 0.070 11, 690
1640mL 167,000 0.080 13, 360
1740mL 167,000 0.088 14, 696
1840mL 167,000 0.096 16, 032
1940mL 167,000 0.105 17, 535
2020mL 167,000 0.112 17, 704
Total heat Load from rock strata = 101, 701 kW.
5.3 Heat Load from Diesel Powered Equipment
Mufulira mine runs several loaders, dump trucks and mobile equipment of various makes and
sizes. The Zambian mine and mineral acts require that a mine should provide 0.05m³/s of
ventilation air per kilowatt of engine power. However, typical ventilation requirements for
acceptable operation of diesel equipment are 0.035m³/s ‘over the engine’ per kilowatt power.
The heat generated by diesel equipment has been calculated based on their total rated power and
amount by which it is used.
41. 32
Table 5.3: Heat generated by diesel equipment
Equipment No. of Units Rated Power (kW) Total Power (kW)
CATR1600G Loaders 6 200 1200
Saandvik LH410 Loaders 1 220 220
AD55 dump trucks 2 304 608
AD45 dump trucks 1 438 438
AD30 dump trucks 1 485 485
Simba 250/1250
Long hole Rigs 2 64 128
Simba S1D
Long hole Rigs 1 55 55
Boomer H281
Development Rig 4 63 252
Atlas Copcos S1D
Development Rig 1 58 58
Normets 5 40 200
Utility Vehicle 4 40 160
Charging Units 3 40 120
TOTAL 31 3 924
Assuming 80% availability 3,139kW is the heat estimated from diesel powered equipment.
42. 33
5.4 Heat Load from Electrical Equipment
The heat from electrical equipment is mainly from the conveyor belts and auxiliary fans.
Auxiliary fans
In calculating the heat from electrical equipment it was assumed that all input power to auxiliary
fan is finally converted into waste heat except for those used to exhaust air from the mine, they
do not affect the ventilating air underground.
Table 5.4a: Fan characteristics
Fan size Make Diameter (mm) Volumetric
flow (m³/s)
Shaft power
(kw)
30’’ fan woods 760 12 45
48’’ fan woods 1220 23 37
Note: A fresh air intake fan at 1420m level has a shaft power of 75kW.
Table 5.4b: Overall total fan power developed
Level Number of 30”
Φ fans
Total power
developed by
30” Φ
fans(kw)
Number of
48” Φ fans
Total power
developed
by 48” Φ
fans(kw)
Overall total
fan power
developed
(kw)
1340 1 45 1 37 82
1357 3 135 1 37 172
1373 2 90 0 90 90
1390 2 90 1 37 127
1407 5 225 1 37 262
1420 1 45 1 75 120
1423 2 90 1 38 127
1440 4 180 0 0 180
Deeps total fan power developed = 1 160kW
43. 34
Conveyor belts
𝑞 𝑐𝑜𝑛𝑣 = 𝐸𝑙𝑒𝑐𝑡𝑟𝑖𝑐𝑎𝑙 𝑝𝑜𝑤𝑒𝑟 𝑐𝑜𝑛𝑠𝑢𝑚𝑒𝑑 − 𝑊𝑜𝑟𝑘 𝑑𝑜𝑛𝑒 (𝑘𝑤)
=
(𝐸𝑙𝑜𝑎𝑑 × 𝑡𝑙𝑜𝑎𝑑 + 𝐸 𝑛𝑖𝑙 × 𝑡 𝑛𝑖𝑙)
(𝑡𝑙𝑜𝑎𝑑 + 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓)
−
𝑀 𝐶 × 𝑔 × (𝑧2 − 𝑧1)
(𝑡𝑙𝑜𝑎𝑑 + 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓) × 3.6 × 106
Where;
𝒒 𝒄𝒐𝒏𝒗 = Conveyor heat load, (kW)
𝑬𝒍𝒐𝒂𝒅 = Electrical power at average belt load (kW)
𝑬 𝒏𝒊𝒍 = Electrical power at belt load (kW)
𝒕 𝒐𝒇𝒇 = Time the conveyor is off, (hours)
𝒕𝒍𝒐𝒂𝒅= Time the conveyor runs loaded, (hours)
𝒕 𝒏𝒊𝒍 = Time the conveyor runs nil loaded, (hours)
𝑴 𝑪 =Mass moved in total time(𝑡𝑙𝑜𝑎𝑑 + 𝑡 𝑛𝑖𝑙 + 𝑡 𝑜𝑓𝑓), Kg
Table 5.4c: Conveyor heat load
Conveyor
Elevation Motors Voltage Full
load
current
Off load
current
𝐸𝑙𝑜𝑎𝑑
(𝑘𝑤)
𝐸 𝑛𝑖𝑙
(𝑘𝑤)
from to
Plan C
Conveyor 1400ml 100ml
15×160kW 550V 207A 80A 2400 660
1×55kW 550V 70A 32A 55 18
Table 5.4d: Mass conveyed
Date Mass conveyed (𝑀 𝐶)
(Tonnes/day)
Mass conveyed (𝑀 𝐶)
(𝑘𝑔/day)
24/08/13 2745 2745000
28/08/13 2205 2205000
29/08/13 2402 2402000
05/08/13 2236 2236000
45. 36
5.5 Heat Load from Humans
The amount of heat produced by the human body varies depending on the amount of work
performed. According to the research carried by C. H. Wyndham in 1971, Human Sciences
Laboratory, Chamber of Mines Research Organization, the following data in Table 5.5a was
provided;
Table 5.5a: Metabolic heat
Amount of work Heat produced (W)
At rest 90
Light rate of work 200
Moderate rate of work 275
Hard rate of work 470
Table 5.5b: Manpower at deeps section
Number of workers at Mufulira Deeps as at 9th September, 2013
Mopani employees contractors Total
88 597 685
Assuming work is between moderate and hard rate of work, then;
𝑞human = (
275+470
2
)𝑤 = 372.5w
Therefore, Total 𝑞human = 685 × 372.5w = 255kw
46. 37
5.6 Total Mine Heat Load
Total mine heat load is the sum of all the heat transfer that take place in the ventilating air
streams.
Table 5.6: Total mine heat load
1440mL 1540mL 1640mL 1740mL 1840mL 1940mL 2020mL
Auto
compression
(kW)
523 1 254 2 165 4 609 6 463 8 596 11 200
Heat from the
rock (kW)
10 354 11 690 13 360 14 696 16 032 17 535 17 704
Heat from diesel
equipment (kW)
3 139 3 139 3 139 3 139 3 139 3 139 3 139
Heat from
electrical
equipment (kW)
2 485 2 485 2 485 2 485 2 485 2 485 2 485
Metabolic Heat
(kW)
255 255 255 255 255 255 255
Other sources of
Heat (kW)
1 862 2 091 2 378 2 798 3 153 3 557 3 865
Total Heat (kW) 18 618 20 914 23 782 27 982 31 527 35 567 38 648
Total mine heat load = 38 648 kW
Total Heat discharge into the mine air stream from Auto compression, the rock, water, men,
mechanical and electrical equipment has carefully been analyzed based on planned production
target of 167, 000 tonnes per month to be 38 648kW for the Deeps, 1340m to 2020m levels.
47. 38
5.7 Air quantity required to dilute the total mine heat load.
Air quantity (Q) required to dilute the total mine heat load is calculated based on the variability
of the working place temperatures. In this study the current and target temperatures were
considered.
Table 5.7: Psychrometric properties
Current working conditions Target working conditions
Temperature = 30/34 °C Temperature = 28/32 °C
Pressure = 118 kPa Pressure = 118 kPa
Density = 1.17 kg/m³ Density = 1.17 kg/m³
Sigma heat = 99.2 kJ/kg Sigma heat = 78.8 kJ/kg
Q =
Total Heat Discharge Into The Air Stream
Average Air Density x Change in Sigma Heat
=
38 648
1.17 x (99.2−78.8)
= 1, 619m³/s
5.8 Calculation for refrigeration ventilation requirement
According to Mike Romaniuk (Hard Rock Miner's Handbook 5th
Edition), a deep hot mine
should resort to refrigeration ventilation if air quantity (Q) required to dilute the total mine heat
is greater than air quantity limit (Qlimit).
Qlimit = (4.74 X 10-4
m³/s.tonne) X (annual ore production)
Qlimit = (4.74 X 10-4
m³/s.tonne) X (2 000 000 ton)
= 948m³/s
Since, Q (1619 m³/s) > Qlimit (948m³/s), Mufulira mine Deeps section should resort to
refrigeration ventilation system.
48. 39
CHAPTER SIX: VENTILATION REQUIREMENTS
6.1 Proposed ventilation circuit
Each production panel is ventilated by a secondary intake ventilation raise from the decline. The
raise connect to each strike drive, and transfer air to the uppermost available crosscut. As mining
advances, the uppermost crosscuts is no longer required as the main traveling way into
production panel, and is converted from an intake to a return airway. On each level a ventilation
duct is installed in the strike drive, with tee-pieces and ducting ventilating each of the strikes.
Either exhaust or forcing system could be used.
Figure 6.1: Proposed ventilation circuit
49. 40
6.2 Proposed main intake airways between 1340m and 2020m levels
When mining takes place on all levels between 1340m and 2020m levels the following intake
airway infrastructure should be available:.
Table 6.2: Main intake airways
Airway
Depth
Dimensions
(m)
Size
(𝑚2
)
Velocity
target
(𝑚/𝑠)
Quantity
flow rate
Target
(𝑚3
/𝑠)From To
56 Ramp
Decline
1340ml 2020ml 5(h) x 6(w) 30 10 300
62 Ramp
Decline
1340ml 2020ml 5(h) x 6(w) 30 10 300
61/62 VR 1340ml 1440ml 4.5(h) X 2.8(w) 12.6 14 176
55P8 VR 1340ml 1640ml 4.5(h) X 2.8(w) 12.6 14 176
56P9 VR 1340ml 1740ml 4.5(h) X 2.8(w) 12.6 14 176
63P3 VR 1340ml 1840ml 4.5(h) X 2.8(w) 12.6 14 176
New shaft
[75%
utilization ]
1340ml 2020ml 6.7(diameter) 35.3 12 318
Total capacity for main intake airways 1 622
6.3 Proposed main return airways between 1340m and 2020m levels
When mining takes place on all levels between 1340m and 2020m levels the following return
airway infrastructure must be available:.
50. 41
Table 6.3: Main return airway raises
6.4 Proposed refrigeration plant
The energy and heat balance done during data analysis (Chapter 5) indicated that a mine
refrigeration plant is required. The suggested refrigeration plant requires an 8 MW air cooling
duty, with this amount of cooling, air will enter the mine at approximately 14 °C.
The refrigeration plant itself can be located at the top of No. 9 shaft.
Figure 6.4: Surface plant and bulk air cooler - Impala platinum mines, South Africa
Raise
Depth
Dimensions
(m)
Size
(𝑚2
)
Velocity
target
(𝑚/𝑠)
Quantity flow
rate target
(𝑚3
/𝑠)
From To
64P5 VR 1340ml 2020ml 6(diameter) 28.3 12 400
58P6 VR 1340ml 2020ml 4(w) X 3(h) 12.0 12 144
59P6 VR 1340ml 2020ml 4(w) X 3(h) 12.0 12 144
13 internal
raises
3(w) X 2(h) 78.0 12 936
Total capacity for main return airways 1 624
51. 42
6.5 Proposed underground fans
In order to optimize the current ventilation system in the deeps section the following booster and
auxiliary fans are proposed.
Booster fans
The existing booster fans used underground must be replaced with high pressure axial fans (fan
performance curve shown in figure 6.5).
Figure 6.5: Booster fan performance curve
Auxiliary fans
High duty auxiliary fan (30 inches/760mm diameter rated at 14.0 m3
/s at 2 kPa) are
recommended for the deeps section.
52. 43
CHAPTER SEVEN: CONCLUSION AND RECOMMENDATIONS
7.0 Conclusion
The following are the findings;
I. There is inadequate ventilation in most working areas of the Deeps section.
II. Air reaches the deeps section at the temperature of 28°C wet bulb, hence little cooling is
done to the working places.
III. There are no Main Return Airways in strategic locations (e.g. at 1340m and 1440m
levels)
IV. There is delay in mining ventilation raises and construction of concrete seal walls
V. There is a lot of Ventilation leakages through old underground workings and worn out
ventilation ducts.
VI. There is too much water in haulages which carries with it a lot of heat which is later
passed on to the air as it moves through the sections.
VII. There is indiscriminate dumping of waste rock in return airways which increase the
resistance to air flow.
VIII. There are poor practices in secondary ventilation. The standard procedures are most often
not followed, for example the section has an average distance of force columns to faces
of 15m instead of 5m. Long and poorly maintained columns are prevalent.
7.2 Recommendations
The following rehabilitation action needs to be implemented in order to improve the existing
ventilation conditions in the Deeps section.
Intake airways
Slyping of intake raises between 1040mL and 1340mL must be completed as soon as
possible.
Holing the second intake RBH between 1430mL and 1423mL must be done and the removal
of loose rock to be done concurrently.
53. 44
Return airways
In order to achieve the required air flow carrying capacities in the Deep section the following
infrastructure is required. The following rehabilitation actions are recommended:
Complete Slyping of the return raises between 1040mL and 1340mL.
Mining of 21 internal raises between 1323mL and 1440mL.
Lashing of blocked return airways on 1340mL at 61 P2 and 62 P8 raise lines.
Outlined below are suggestions to achieve the ideal ventilation system for Mufulira mine Deeps
section in order to provide a safe environment both for personnel and equipment.
I. An 8 MW refrigeration plant should be installed to reduce the intake air temperature into
the mine and this can be located on the top of No.9 shaft.
II. A third fan should be installed at number 10 shaft. This fan will provide an additional
exhaust capacity of 212m³/s.
III. The Deeps section air flow is dependent on the slyping of the internal return raises
between 1040m and 1340m levels. This project is currently on-going and must be
completed as soon as possible.
IV. Fan requirement, six thirty inches force fans and four forty-eight inches exhaust fans per
level i.e. a set of three thirty inches force fans and two forty-eight inches exhaust fans on
the western and eastern sides each.
V. Return airways should be mined ahead of production operations and construction of
concrete seal walls should be done within the shortest period of time.
VI. Stopped out areas must be sealed using brick walls to avoid dilution of fresh air.
VII. Worn out ventilation ducts must be replaced with good ones to avoid air leakages.
VIII. Ventilation raises must only be used for ventilation not for tipping of ore and waste
which mostly lead to choking of ventilation raises.
IX. All the water from underground must be directed into water drive instead of footwall
drive.
X. Ventilation department should purchase the Vuma-3D ventilation software in order to
properly manage environmental conditions encountered underground.
54. 45
REFERENCES
1. Anon., 1971, “Ventilation Planning as a Prerequisite for Winning Higher Outputs”, Mining
Engineer, Vol. 30, Part 12, pp. 796–811
2. Burrows, J., et al., eds., 1982, “Environmental Engineering in South African Mines”, Mine
Ventilation Society of South Africa, Johannesburg.
3. Chambers of Mines of South Africa, “Measurements in Mine Environmental Control”,
Johannesburg, 1982.
4. Geology, planning and Ventilation Departments MCM Mufulira mine.
5. Hartman, H.L., Mutmansky, J.M., and Wang, Y.J., eds., 1982, “Mine Ventilation and Air
Conditioning, 2nd
ed.”, Wiley Interscience, New York.
6. Hoffman, D. and Marx W., 2013, “Ventilation and cooling study for Mindola and Mufulira
shafts”, BBE Consulting, South Africa.
7. Jack de le vergne, 2003, “Hard Rock Miner's Handbook Edition 5”, Arizona, USA.
8. Le Roux, W. L., 1979, “Mine Ventilation Notes for Beginners, 3rd
edition”, The Mine
Ventilation Society of South Africa.
9. Marx WM et al, (2008) “Development of Energy Efficient Mine Ventilation and Cooling
Systems”, Mine Ventilation of South Africa Society Journal, April/June.
10. Mining Education Australia, 2006, ‘Underground Mine Environment Course Modules 3.0
Heat in underground mines and 7.0 Metal mine hazards and control’.
11. O’Neil, T.J., and Johnson, B.R., 1982, “Metal Mine Ventilation Systems,” Mine Ventilation
and Air Conditioning, 2nd
edition, Chap. 14.
12. Roberts, A, 1960, “Mine Ventilation, Ventilation planning Chap. 13 P260-261”, Cleaver
Hume Press Ltd., London.
13. Zambia Mining Regulations act of 1974, ‘A Guide to Mining Regulations, Part (ix)’.
56. 47
APPENDIX II
Acceptable environment standards
The following are the general acceptable ranges of air velocities in Primary airways that could
create acceptable environmental conditions in the mine
Description of Area Velocity (m/s)
Development ends / Draw points 0.5 – 4.0
Ramp Raises 2.5 – 8.0
Conveyor Tunnels 2.5 – 8.0
Main Haulages 2.5 – 8.0
Main Shaft Crosscut 4.0 – 8.0
Down Cast Shafts 8.0 – 10.0
Up Cast Shafts (Concrete lined) 1.5 – 20.0
Up Cast Shafts (Rough Rock) 10.0 – 15.0
Return Ventilation Raises (Concrete lined) 15.0 – 20.0
Return Ventilation Raises (Rough Rock) 10.0 – 15.0
Optimum wet bulb temperature for Zambian mines = 31.0 °C
57. 48
APPENDIX III
Deeps Section Ventilation Status
Intake Work Place
The temperature = 26.0/28.0o
c The temperature = 30.0/34.0o
c
Dust concentrations = 40ppcc Dust concentrations = 90ppcc
Relatively Humidity = 70% Relatively Humidity = 65%
Maximum Allowable dust concentration = 120ppcc
Toxic gases present in the atmosphere
Gases CO CO2 NOX
Average 13ppm 1300ppm 3ppm
Maximum Allowable 1000ppm 7500ppm 10ppm
Toxic gases present at the exhaust
Gas CO NOX
Average 4000ppm 3000ppm
Maximum Allowable 2000ppm 1000ppm
