Document Sample


             MIN - 03018

       PART - I

           B.Tech. First Year

          Mining Engineering


                                            CHAPTER 1
1.1       General Information and Hazards
          Oxygen is colorless, odourless and tasteless and slightly soluble in water. The main
hazard of oxygen is not poisoning but by the lack of it, i.e. oxygen deficiency.

      Oxygen %                     Effects
      21                           Normal concentration
      17                           Breath faster and deeper
      14                           Dizzy sensation, buzzing in the ear
      10                           Increased heartbeat
      7                            Confusion, loss of consciousness and death

          Nitrogen is colorless, odourless and tasteless and practically insoluble in water. It is
chemically very inactive and its main hazard is that it dilutes the oxygen in the atmosphere,
i.e. causes oxygen deficiency.

Carbon dioxide
          Carbon dioxide (CO2) is colorless, odourless and has a faint acid taste. It is denser
than air and does not burn or support combustion. Its main hazard is that it dilutes oxygen in
the atmosphere, i.e. causes oxygen deficiency. It also acts as a stimulus to the breathing rate
of a person, as indicated in the following table.

      Carbon dioxide %             Increase in respiration rate
      0.5                          Slight
      2.0                          50%
      3.0                          100%
      5.0                          300%
      10.0                         Intolerable for extended periods

       First aid measures are to remove the patient to fresh air, give oxygen or artificial
Carbon monoxide
       Carbon monoxide (CO) is colorless, odourless and tasteless. It does not support
combustion but will burn and explode at concentration between 12.5 and 74%.
Preferentially fastens to haemoglobin in the blood in place of oxygen. Brain and heart
muscles, the greatest oxygen consumers are most affected. Above 4000 ppm, coma can
occur without warning. 500-1000 ppm can cause headache, first breathing, nausea,
weakness, dizziness, mental confusion and hallucination.

Hydrogen sulfide
       It can be produced if acid contacts metal sulfides. And if traces of arsenic and
phosphorus are present, arsenic and phosphene can also be generated, both of which are
toxic. It is more toxic than hydrogen cyanide. It irritates eyes and respiratory tract at low
concentrations. Higher concentrations cause respiratory paralysis with asphyxia,
convulsions and coma.

Description: Silvery liquid; metallic element (Hg)
Constants: At.wt. = 200.61; mp = 38.89° C; bp = 356.9° C; d = 13.546.
TLV : 0.1 milligram per cubic meter of air (ACGIH)
Hazard: Dangerous; when heated it emits highly toxic fumes
Countermeasures: Ventilation control, storage and handling, and shipping regulations.
       Mercury has disastrous results when inhaled as gas. In the lungs, it slowly becomes
soluble as methyl mercury. It is then absorbed into the blood stream, penetrates the blood/
brain barrier, and causes acute mental and muscle decline.
       After absorption, it circulates in the blood and stored in the liver, kidneys, spleen
and bone. It is eliminated in the urine, feces, sweat, saliva and milk. In industrial poisoning,
the chief effect is on the central nervous system, mouth and gum. Colitis has been reported
frequently and nephritis is rarely reported.
       Other symptoms of mercury poisoning are stomatitis, tremors, and psychic
disturbances. The first complaints are of excessive salivation and pain on chewing. In severe
cases, there may be loosening of teeth, and a dark line on the gum margins. In slow
poisoning, the salivation may be absent, and the only complaint is dryness of the throat and

mouth. The tremor is of intention type. It can be seen when the patient spread the
outstretched fingers, protrudes tongue, or attempts to perform specified movements.
Muscles of the face, hands and arms are chiefly affected. In more severe cases, there may be
convulsive movements. Writing is illegible. The psychic disturbances include loss of
memory, insomnia, lack of confidence, irritability, vague fear and depression.
       The dermatitis produced by fulminate of mercury takes the form of small, discrete
ulcers on the exposed parts, and usually accompanied by inflammation of mucous
membranes of the nose and throat.

Cyanide Compounds
       Cyanogens (CN)
       m.w                     52.04                   0.87
       m.p                     - 27.9 C
       b.p                     - 20.7C
       Very soluble in water, ethyl alcohol, ethyl ether. A colorless gas with an almond-like
ordour, which is crid and pungent in lethal concentrations.
       TLV                     (ACGIH) 10 ppm 20 mg/ m3
       Acute exposure can cause death by asphyxia. Chronic exposure to cyanides at levels
too low to produce serious chemical complaints are known to cause a variety of problems,
Study of workers in the electroplating industry has shown dermatitis to be a problem there.
Also reported were itching, scarlet rash, papules, in addition to severe irritation of the nose,
leading to obstruction, bleeding, sloughs and in some cases perforation of the septum.
Among fumigators, mild cyanide poisoning is recognized as the cause of the symptom of
oxygen starvation, headache, rapid heart rate, nausea, all of which are completely reversed
when exposure ceases.

       Acute exposure can cause death by asphyxia. Chronic exposure to cyanides at levels
too low to produce serious chemical complaints are known to cause a variety of problems,
Study of workers in the electroplating industry has shown dermatitis to be a problem there.
Also reported were itching, scarlet rash, papules, in addition to severe irritation of the nose,
leading to obstruction, bleeding, sloughs and in some cases perforation of the septum.

Among fumigators, mild cyanide poisoning is recognized as the cause of the symptom of
oxygen starvation, headache, rapid heart rate, nausea, all of which are completely reversed
when exposure ceases.
               There are a series of cases in the literature indicating that chronic systematic
cyanide poisoning may occur but this is rarely recognized because of the gradual onset of
disability and the appearance of symptoms which are consistent with other diagnoses.
Excessive thiocyanate in extra-cellular fluids might explain chronic illness due to cyanide,
as the symptoms reported are similar to those found when thiocyanate is used as a drug. The
occurrence of symptoms of chronic disease have been reported in electroplaters and silver
polishers after several years of exposure. The most prominent effects were motar weakness
of arms and findings have also been reported as complications of thiocyanatic therapy.
       In spite of the extreme hazard of cyanide exposure, there are far fewer industrial
cases of acute poisoning than might be expected. Furthermore, where emergency measures
were well understood and contingency plans carefully developed, seriously poisoned
workers have been successfully treated.

       The cyanide ion is rapidly absorbed from all routes of entries, respiratory and
gastrointestinal tracts, as well as through unbroken skin. The toxic properties of cyanide
depend on its ability to inhibit enzymes required for the respiration of cells. Thus it prevents
the uptake of oxygen by the tissues and causing death by asphyxia. In poisoning, the
cyanide gets into the blood and blocks oxygen transfer from the blood to the body tissues.
There is oxygen in the block, but it is tied up and can't get to the vital organs, This is known
by the pink or even brick-red skin color of the victim. The effect of this oxygen blockage is
similar to drowning, a reduced supply of oxygen to the vital organs.

Storage and Handling
       The following should be considered in storing and handling cyanide.
       1. Cyanide salts should be stored in dry ventilated enclosure. This enclosure should
           be lockable and appropriately signed with regard to the toxic hazards and safety
           precautions applicable to the contents.
       2. Storage areas should be kept clean. Vacuuming is the preferred method for dry
           cleanup. Water should be used only after the surface has been thoroughly

   vacuumed or swept. Storage area floor drainage should be connected to the
   cyanide circuit to facilitate the safe disposal of liquid spills.
3. Cyanide salts should not be stored where they can be exposed to acid vapors or
   oxidizing agents (hypo-chlorides, peroxides or chlorates).
4. Cyanide salt should be kept in the original shipping container until used. Large
   shipping containers loaded with smaller cyanide drums or bags should be
   approached with caution. Rough handling of the container can result in drum or
   bag bursting with subsequent contamination of the entire container with
   hydrogen cyanide. HCN is a colorless gas with an almond like odor. It can be
   measured utilizing stain tubes or handheld direct reading monitors. Definitive
   laboratory methods, which can also be utilized in the field, are also available.
   Tests should be carried out prior to entering any confined, non-ventilated space
   suspected of cyanide contamination.
5. The consumption of food, drink and tobacco should be prohibited in cyanide
   storage and mixing areas. Employees should be encouraged to maintain the
   highest standards of personal hygiene. Cyanide should not be stored or handled
   near any food storage, preparation or dispensing area. Employees should remove
   contaminated clothing and washed thoroughly before entering lunchrooms.
   Contaminated clothing should be safely laundered commercially. Lunchrooms
   should be kept clean and uncontaminated.
6. Cyanide mixing: The mixing of cyanide salts should be conducted in such a
   manner that the emission of dust is minimized. Drum handling should be
   automated where possible and enclosure of the entire mixing process is
   recommended. If cyanide drums have to be opened manually, special tools
   should be used rather than the ubiquitous cold chisel. The pH of the mixing
   solution should be maintained at pH 10 to preclude the possibility of hydrogen
   cyanide generation.
7. In situ leaching: All accessible underground workings used for the collection of
   cyanide leachates should be provided with positive mechanical ventilation. All
   cyanide circuit components should be clearly labeled.
8. Carbon dioxide fire extinguishers should not be used in cyanide areas. Carbon
   dioxide is mildly acidic and will thus produce hydrogen cyanide on contact with
   cyanide salts. Cyanide salts will not burn or support combustion. If chlorination
   is used for the treatment of cyanide wastes particular care should be taken when

          storing calcium hypo-chloride. This compound is fairly unstable and sharp
          knocks can cause ignition. Avoid contact with organic materials and utilize
          copious quantities of water top clean up spillages.

Environmental Control in Cyanide Operations
       Cyanide is not only very toxic but also polluting to the environment and this hazard
       will be controlled by prevention rather than cure. Cyanide operations should be well
       planned and engineered to achieve close control of the following sources of cyanide
       1. Seepage from tailing dams, leach pads etc.
          Although measures for controlling seepage may be found acceptable on some
          sites, there will be more ready approval for tailings dams and leach pads to be
          lined and/ or dewatered.
       2. Run-off from contaminated areas
          High standards of house keeping are expected on the sites of cyanide operations
          but some spillage is inevitable. Areas of catchment which are likely to become
          contaminated should be incorporated in closed water management systems.
       3. Overtopping of dams
          There is always some risk that the design capacity of a dam may be inadequate
          during a period of high rainfall and high run-off. Risks should be calculated and
          will be considered prior to approval.
       4. Escape of cyanide after operations have ceased
          Cyanide operations should be safely decommissioned so that control persists as
          long as the cyanide is potentially toxic. The assertion of the industry that residual
          cyanide in the tailings is rapidly transformed into innocuous chemicals needs to
          be proved on a case by case.
          Residual cyanide in leach dumps requires attention. If the quality is unacceptable
          for discharge, the spent ore should remain on its pad and the closed circuit, nil
          release, retention system should continue to operate. If the leachate is acceptable
          for discharge, opportunities for reshaping spent ore beyond the pad arise.

Disposal of Cyanide
       In general, the most convenient way of disposing cyanide waste is by oxidation at a
pH of 8.5. The cyanide is rapidly converted to cyanate by an oxidizing agent. Cyanates can

be eliminated by further chlorination. Other products such as "Cyanide hydratese" produced
by ICI can be used to detoxify cyanide waste waters.

Personal Protection
       The cyanide mixing area should be provided with the following:
       (a) Safety shower
       (b) Eye wash
       (c) Hand washing facilities
       (d) Wash down water hose
       (e) Adequate ventilation to ensure contaminated air is ducted away from the area.
       Personnel engaged in handling cyanide salts should wear the following equipment:
       (a) Protective eye goggles
       (b) Appropriate respiratory protection
       (c) Waterproof clothing, face shields and gloves
       (d) Rubber boots or overshoes
       All the above equipment should be subject to a regular cleaning and maintenance
program. Protective equipment suitable for use in cyanide environments is available from
most reputable manufacturers.

Rescue Operation
       The victim may know he is in trouble. He will frequently run or stumble a few steps
to get away then fall as his oxygen is depleted. The following rescue procedure is
recommended to help the victim.
       1.   Check to make sure if it is safe to enter the area where the victim is. Stay
            upward. Use an air mask, if needed for rescue, or simply hold your breath for
            short exposure.
       2.   Walk, carry or drag the patient to fresh air.
       3.   Remove contaminated clothing and or wash the patient.
       4.   Have the patient lie down and keep warm.
       5.   Based on the patient's condition, give first aid.
       6.   Get medical help to administer medical treatment.

First Aid and Emergency Equipment
       First aid and medical treatment techniques provide antidotes designed to react with
the cyanide in the blood and detoxify the victim. The toxic effects of cyanide are rapid. But
treatment is effective when provided quickly and recovery is normally rapid from a non-
fatal dose. Unlike many poisons, cyanide is not cumulative, so it does not build up in the
body with prolonged exposure.
       First aid should be given quickly. This requires both equipment availability in the
cyanide use area and people trained to use the equipment. It also requires pre-plan so that
people know what to do and act in an emergency.
       Recommended basic emergency equipment is:
       1. Amyl nitrate antidote
       2. Medical supplies
       3. Oxygen resuscitators
       4. HCN detector
       5. Self-contained air mask.
       The procedure is to break the glass pearl in a cloth and hold under the nose for 15
seconds, and then take it away for 15 seconds. Repeat this "hold and away" procedure five
or six times with each pearl. If necessary, use three to four pearls at a rate of one every three
minutes. The rescuer should avoid amyl nitrate inhalation and should not smoke. Amyl
nitrate must be replaced every one or two years, as aging can pressurize and break the glass
pearls. All of this equipment should be available where cyanide is used and inspected
regularly by the people who will be using it.
       In the case of cyanide swallowing, have the patient drink one or two pints of sodium
thio-sulfate or plain water, then induce vomiting with finger in throat. Repeat until vomiting
is clear. Remember not to give anything to someone unconscious. Two one-pint bottles of
one percent sodium thio-sulfate are recommended as first aid supplies.

Training Requirements
       Prevention is the bvest way to avoid accidents. Workers should be trained so that
they know how to store and handle cyanide and cyanide solutions, how to use their personal
safety equipment.

1.2    Sources
1.2.1 Strata gases
       These occur in the mineral deposit itself such as methane in coal measures or
adjacent formations. Migration to the mineral deposit can result from the release of pressure
with mining, carried with fissure water or penetration through porous rock. Strata gases may
be released gradually or more rapidly from blowers. They are not confined to sedimentary
rocks and occur in mines with igneous rock structures. the main strata gases are methane,
carbon dioxide, nitrogen dioxide, nitrogen, sulphur dioxide, hydrogen sulphide and radon.

1.2.2 Explosives
       The amount and type of gases produced by the detonation of explosives depends on
the type of explosive being used. The following table illustrates the results of tests on
different explosives under varying conditions.
Table Gases Produced by Explosives

 Type of explosive                   CO             CO2            NH3           NOx
 Ammo-gelinite (hard rock)           0.037          0.153          0.008         0.012
 Ammo-dynamite (hard rock)           0.025          0.117          0.006         0.010
 Ammo-dynamite (soft rock)           0.019          0.117          0.012         0.015
 Ammo-dynamite (water in hole)       0.013          0.091          0.009         0.014
 ANFO (hard rock)                    0.020                                       0.020
 ANFO (soft rock)                    0.010                                       0.040

1.2.3 Diesel equipment
       Petrol driven engines are not normally used underground because of the large
quantities of carbon monoxide produced. Diesel engines are frequently used and they can
produce up to 0.2 m3 /s of exhaust gases per 100kW rated diesel power. The amount and
chemical composition of the exhaust gases will depend on the following:
       1. Working load of the diesel, idling to full load
       2. Type of engine including manufacturing differences
       3. Maintenance of the engine and adjustment
       4. Type of fuel being used

       It is therefore difficult to provide absolute value of the amount and composition o
exhaust emissions. Typically, the toxic exhaust gases are carbon dioxide, carbon monoxide,
oxides of nitrogen and hydrocarbons. If the fuel contains sulphur, then some sulphur dioxide
will also be present in the exhaust gases.
Oxidation, fires and other equipment
       The gases resulting from oxidation, fires, explosions and other equipment are mainly
carbon monoxide and carbon dioxide. Acetylene from welding and cutting equipment,
hydrogen from battery charging and other gases resulting from burning plastic materials can
also occur depending on the situation.

1.3    Control
1.3.1 Prevention of Formation
       The prevention of formation of the gas is not only a very simple approach but also
usually the most cost beneficial. The correct adjustment and maintenance of diesel
equipment is an important means. Poor maintenance increases the concentrations of the
toxic components of the exhaust gases. This is readily detected by an increase in
hydrocarbons and in particular aldehydes which irritates eyes, nose and throat. Most diesels
used in underground are derated to 85% of their surface output and have some form of
exhaust cleaning equipment.

1.3.2 Removal
       Many coal mines have methane drainage systems. In some mines, methane is piped
to surface and used to fire a boiler. More generally, removal is achieved by local exhaust
systems. It is important to recognize the removal by itself is merely transferring the hazard
from one location to another and some additional control method may be required.

1.3.3 Isolation
       Where the gases cannot be avoided from forming or removed by local exhaust
systems, it may be possible to locate the source to ensure that personnel are not exposed.
One example of this is during blasting. Normally blasting only takes place when personnel
are removed to surface and an adequate time allowed for the gases to disperse or be
removed before men return underground. A second example is the sealing off of worked out
areas in a mine here radon emanates from rock surfaces. This reduces the emanation surface

area but may result in secondary problems when high concentrations behind sealed areas
contaminate working areas.
1.3.4 Dilution
       Dilution of gaseous contaminants to reduce them to acceptable concentrations is the
most widely used and most successful method of control. The basic dilution formula is:

       Q = ---------------- - Qg
             AC - B

       Where Q = the required dilution rate m3 / s
              Qg = gas emission rate m3 / s
              AC = allowable gas concentration
              B = concentration present in normal air

1.3.5 Legislation
       Legislation is broadly classified as statutory or discretionary. The problem is that the
 limit by itself does not satisfy the objective of legislation. This is to ensure that safe
 working conditions exist in the underground environment. This can only be achieved by
 the continuous monitoring of all working places which is clearly impractical. Spot samples
 are therefore the norm.
       Legislation is therefore difficult because in some cases the limit may be too severe
and in others too lenient. Statutory limits are those actually written in a Mines and Work
Act. Discretionary limits often occur in codes of practice where the local mining
inspectorate can impose more stringent limits than may be laid down in statutes.
       In 1910, the US Bureau of Mines was formed which set standards but had no
enforcement power. It was not until 1966 (metal mines) and 1969 (coal mines) that Federal
standards became enforceable. The Federal standards are based on the threshold limit values
published by the American Conference Of Governmental Industrial Hygienists. Individual
states can require more stringent requirements. Some examples of Federal standards are:
       Carbon dioxide                 0.5% (5000 ppm)
       Carbon monoxide                50 ppm (could change to 35)
       Nitric oxide                   25 ppm
       Nitrogen oxide                 5 ppm
       Oxygen (minimum)               19.5 %

                                         CHAPTER 2
2.1    General Hazards
       The chronic or long term effects of dust inhalation are not limited to the respiratory
system and in many cases the respiratory system provides a mode of entry into the body.
       The respiratory system is also selective with respect to the proportion of dust
retained in the system and also the site of damage. There is no simple one to one
relationship between the concentrations of an atmospheric contaminant, the duration of
exposure and the rate of dosage to the critical site in the body.
In many cases, allowable concentrations are subject to differences of opinion in the
interpretation of data. Additional problems are that the subject can be exposed to several
contaminants at the same time in which case it is extremely difficult to identify and quantify
the separate effects. Further, the social habits of the subject may vary i.e. smokers and non-

2.2    Biological Effects of Dust
       The function of the lung is to supply oxygen to a membrane where it can diffuse
through and be taken up by the blood stream. In return, carbon dioxide is passed from the
blood steam through the membrane into the lungs from which it is exhaled.
       Excessive inhalation of dust particles would result in the accumulation in critical
areas of the lungs. No permanent damage is done and there is no disability. This is
sometimes known as benign pneumoconiosis or nonfibrogenic pneumoconiosis.
       Some dusts such as silica decrease the active life of the cell in the lung, to hours or
days. These dusts are not removed from the cell, and a less controlled accumulation of dust
occurs. The net result is that because the dust is not removed, the tissues become
permanently damaged (scarred) and there is permanent alteration of the cell structures. The
fact that the tissue becomes scarred leads to the term fibrogenic pneumoconiosis.
       Of particular importance is the pneumoconiosis resulting from silica dust, known as
silicosis. The disease is fibrogenic (scar tissue) and the loss of lung capacity leads to
shortness of breath and a lessened capacity for work. The damage in the lung also results in
an increased susceptibility to tuberculosis. The disease is caused by free crystalline silica
rather than silicates. The dust of naturally occurring silica and freshly fractured silica is
more dangerous than old dust.

2.3    Sources
       The main source of dust in mines stems from the rock itself. The mining operations
consist of breaking the rock in situ to sizes convenient for transportation. It is inevitable that
large quantities of dust will be produced. One cubic millimeter of rock crushed to two-
micron particles would yield some 200 million particles. Drilling a hole 30 mm in diameter
and 1.8 m deep has a volume of over 1 million cubic millimeters. Fortunately most of the
dust particles produced when drilling are much larger than two micron and very few of them
become airborne. Generally this is true of most sources of dust if suitable control methods
are applied.

2.3.1 Blasting
       Blasting produces very large quantities of dust. Some of the fine dust is carried away
by the air stream but a large amount of dust is trapped within the broken rock. Large
particles tend to settle out as do fine particles which agglomerates into lager masses. The
amount of dust formed by a blast depends to a certain extent on the pattern of the round, the
method of blasting, the type of tamping material and the type of explosive used.

2.3.2 Mechanical Loading
       The blasted rock contains dust particles both within the rock pile and also on the
rock surfaces. Mechanical loading of the rock will release all the dust contained within the
rock pile and some of the dust on the rock surfaces. The greater the agitation of the broken
rock, the greater will be the amount of dust released. It has been demonstrated that the rate
of dust production is directly related to the hoe in scraping. One of the problems with
mobile loading equipment is ensuring that the exhaust is not directed onto rock surfaces or
the footwall and so stir up the settled dust. The advent of high capacity Load Haul Dump
equipment has created additional problems because their rate of removing broken rock
exceeds most dust allaying procedure.

2.3.3 Drilling
       When introducing the sources of dust in mining, mention was made of the possible
dust production when drilling. Supplying water at the correct pressure and flow rate ensures
that the rock surface is wet at all times, and in fact the rock is broken under a film of water.
Leakage of compressed air into the water can however break down this film and allow dusty
air to escape from the hole. Also the compressed air may assist in evaporating water which

in its own right contains large amount of dust. Water is normally supplied to the drill steel
via a water tube in the drill itself. This water is delivered to the bottom of the hole through
the drill steel. It is generally not very effective when collaring the hole.

2.3.4 Transportation
        Whenever rock changes direction it is normally subject to an acceleration followed
by a rapid deceleration. Dust particles adhering to the surface of the rock may then be
dislodged and become airborne. Typical situations are conveyor transfer points, loading
trucks, discharging trucks, etc.

2.3.5 Miscellaneous Sources
        Diesel engines produce particulates in addition to gases in their exhausts. The
concentrations depend on the mode of operation of the engine and can vary between 25 and
250 mg/ m3 when idling and at full load respectively. Typical average emissions are of 100
mg/ m3. The particulate formed is very fine, probably less than 0.5 microns in size, which
agglomerates into particles with a mean size of between a quarter and half a micron.
                Other miscellaneous sources of dust are blowing out with compressed air or
the misuse of compressed air in general. Welding can produce copious amounts of fine dust.
Most other sources are the results of making airborne previously settled dusts. Intake shafts
and fans close to cement storage silos and dry tailings dams can result in high intake dust

2.4 Control
        The most important method of suppression and control of dust is to avoid creating
them. This simple approach can usually result in maximum benefits at minimum cost and
can be termed source control. If a dust source cannot be eliminated, then the second method
of control is to isolate it i.e. remove the personnel or alternatively remove the dust as it is
formed. A third method is to dilute it as it is formed, with fresh air. If the dust source is very
large, it is often desirable and necessary to filter it in order to prevent its dispersion to the
general atmosphere.

2.4.1 Source Control
        When dealing with dust formed as a result of rock breaking, source control is best
achieved by the application of water. The basic principle is to prevent the dust becoming

airborne. It is important to recognize that the water forces (quantity and pressure) must be in
equilibrium with the forces creating the dust at the point of impact. If they are not, the
application of water will not be fully effective or there will be a wastage of water.
Maintaining a moisture content of 1 % particularly when transporting broken rock through
airways, consequently a 5 % moisture content should be the target value.
        The excessive use of water can result in wastage, flooding of drainage and of ore
passes, overloading pump stations and dirty water settlers. To prevent wastage when
wetting down, a suitable spray nozzle should be used. This should ensure an adequate
spread of water over a large area and prevent the settled dust being stirred up. The water
should be as clean as possible to prevent the formation of dust from the evaporation of dirty
water. Clean water contains up to 10 million particles of dust per milliliter and dirty water
can contain ten or twenty times as much. The use of this water can add to the general
dustiness of the air.

2.4.2 Isolation.
        Where source control is inadequate with respect to high dust levels, isolation is one
of the two back up measures. The removal of personnel from the source is the simplest
method but not always practical. Removal of the dust means the use of local exhaust

2.4.3 Dilution
        In many situations, the dust produced at the source is not of a high concentration and
there may be numerous sources. In these situations, the supply of fresh air to dilute the dust
is advocated. the quantity of fresh air required can be obtained from the dilution formula
given Chapter 2:
        Q = ---------------- - Qg
                AC - B

                Where Q = the required dilution rate m3 / s
                         Qg = gas emission rate m3 / s
                         AC = allowable gas concentration
                         B = concentration present in normal air

        In this case, the rate of formation of the dust and its concentration must be known.
There are four limiting factors to the use of dilution as a method of dust control.
       1.   The quantity of dust generated must not be excessive or the quantity of fresh air
            for dilution will be impractically large.
       2.   Workers must be far enough away from the dust source to allow the dilution to
            take place before they are exposed to the mixture.
       3.   The health hazard of the dust should be low otherwise again the quantity of
            fresh air for dilution becomes impractically large.
       4.   The formation of dust should be uniform and not sporadic.
       In many underground situations, dilution is the only practical method of controlling
the levels of dust encountered.

2.4.4 Dust Collection
       It is not always possible to discharge dust from a local exhaust system into the
atmosphere because of the nuisance value or the possible damage to amenities. The dust
may also be valuable and its losses economically undesirable. The following factors should
be taken into account when considering the selection of dust collection equipment.
       1.   Type of dust: This affects the collection process, solubility, abrasiveness and
            corrosion characteristics.
       2.   Size distribution: This is probably the most important factor since collectors
            only operate efficiently in certain size ranges.
       3.   Required efficiency: From the knowledge of the size distribution of the
            materials to be collected and the efficiency of the collecting method, a decision
            can be made. There is no point in using a settling chamber to collect sub-
            micron particles and it is wasteful to use a fabric dust collector to filter coarse
            dust over 200 microns in size.
       4.   Loading: This is the amount of dust to be filtered per unit of time. If the loading
            is very high the overall efficiency of most collectors decreases. A pre-collector
            could then be considered. Fluctuating loads may also create problems for some
            collection devices.
       5.   Space limitations: This is most pertinent underground where excavation costs
            are very high. Cyclones and wet scrubbers take up very little floor space when
            compared to fabric filters.

        6.   Gas temperature: Fabric filters are not normally used for temperatures in excess
             of 230°C and wet scrubbers may require after cooling too eliminate the steam
             produced from hot processes.
        7.   Removal of the collected material: If the material is valuable and to be returned
             to the process, consideration should be given as to which form it should be
             returned. For example a wet scrubber yields a slurry which would have to be
             dried at extra cost if the material was to be returned to the process in dry form.
             The handling of the material if it is to be disposed should also be examined.
        8.   Maintenance: Estimates of replacement items as well as routine maintenance
             should also be taken into account. In general, the more complex the system, the
             higher will be the maintenance cost.
        9.   Operating costs: Collection efficiencies are generally related to power input.
             The operating costs should be assessed according to the efficiency required and
             should include auxiliary equipment such as fans, pumps and compressors.
        10. Capital costs: It is the initial cost of the equipment including erection costs and
             excavation costs if possible.

2.4.5 Legislation
        In the western world, there are very few statutory limits imposed with respect to dust
levels in metal mines. In the USA, the Federal limits are the threshold limit values proposed
by the ACGIH. Most other countries prefer to use the general statement that dust should be
controlled so that it does not present a hazard to the health of workers. The mine
inspectorates then impose their own discretionary limits. Most countries adopt the limits for
silica given in equations:
Acceptable level = ------------------------ p/ml
                     % SiO2 + 10

Acceptable level = ------------------------ mg/m3
                     % SiO2 + 3

Acceptable level = ---------------------------------------------- mg/m3
                     % respirable fraction SiO2 + 2

                                            CHAPTER 3
3.1    Noise
       Noise can be defined as unwanted sound which is a variation of pressure in the
region of air adjacent to the ear. This variation of pressure is a succession of traveling
pressure waves moving away from a source such as a vibrating surface.

3.2    Velocity of Sound
       The rate at which the successive layers of air are compressed depends on the
elasticity and density of the air. At 20°C it would take one second to compress the air 344
meters along a tube. This gives the velocity of sound in air at 20°C. The process can take
place in any medium that is elastic such as gases, liquids and most solids. Typical velocities
of sound in various materials are given in Table 3.1.

                       Table 3.1. Velocities of Sound in Various Materials

                        Material         Approximate velocity (c) m/s
                        Air              344
                        Lead             1220
                        Water            1410
                        Concrete         3400
                        Wood             3400
                        Rock             5000
                        Steel            5200

       An example of the different velocities of sound in air and in rock is when blasting
takes place underground. The sharp cracks of the explosive being detonated is transmitted
through the rock and can be heard before the long drawn out boom of sound waves traveling
through the air.
       The frequency, wavelength and velocity are related as follows.

                         λ x f = c
       where       λ     = wavelength (m)
                   f     = frequency (Hz)
                   c     = velocity (m/s)

3.3 Sound Intensity and Power
       Sound was defined, and illustrated to be, a succession of traveling pressure waves
moving away from a source. Because they are traveling, the pressure is being transmitted
from the source. This is the same as saying that force is being transmitted or moving.
       Intensity I is the amount of energy passing through unit area in unit time and is
usually expressed as Watts/ m2. Intensity is directly proportional to the square of the
pressure. Since sound measuring instruments measure pressure it is possible to calculate the
total acoustic power produced by a source. This is important because the sound or acoustic
pressures measured depend on the source and its surroundings. Moving the same source into
different surroundings and measuring the pressure at the same distance need not result in the
same values.
       The intensity at a point can be obtained by:

                      I = ---------

               where I = intensity (W/ m2)
                       p = pressure (Pa)
                      w = density (kg/ m3)
                      c = velocity of sound (m/ s)
3.4 Decibels
       The ear is remarkably sensitive and responds to intensity i.e. pressure squared rather
than pressure. The lowest intensity the ear can normally detect is approximately 10 -12 watts/
m2. The threshold of pain is approximately 10 watts/ m2. This very large spread is
compressed into a logarithmic scale and related to a datum value to avoid negative values
and to be more meaningful. If the datum is taken as the threshold of hearing i.e. 2 x 10 -5 Pa
the unit of sound becomes

               log --------------
                    (2 x 10-5)2
       This is known as a Bel and a further modification made to avoid excessive use of
decimals is to multiply it by 10. This then results in the decimal or dB. With respective to
acoustic pressure the units becomes:

Sound Pressure Level SPL = 10 log -------------- dB
                                   (2 x 10-5)2

With respect to acoustic power the reference level is 10-12 watts and the units become:

 Sound Power Level SPL = 10 log -------------- dB
                                   10 -12

3.5     Hearing Damage and Acceptable Levels
        Hearing damage results from damage to the hear cells caused by excessive
vibrations. This is normally progressive and can take place over extended periods of time
i.e. 10 to 15 years.
        Acceptable levels depend on the criteria being used for assessment. From the point
of view of the underground ventilation engineer the usual criteria is the avoidance of
hearing damage. The following table gives two different permissible noise exposures, both
of which are recognized by various international bodies.

        Table 3.2. Permissible Noise Exposures

            Duration in hours per day                 Noise level in dBa

                            8                         90                85
                            4                         95                88
                            2                        100                91
                            1                        105                94
                           ½                         110                97
                           ¼                         115               100

A second criterion that may occur underground is speech and signal interference. This could
occur at telephones, shaft signals at stations or vehicle warning sirens at ventilation doors.

3.6 Typical Noise Levels
       Typical sound pressure levels for various equipment are given in Table 3.2. Note
that because sound pressure levels are given the location of the person affected is also

Table 3.2. Typical Sound Pressure Levels
Sound Pressure Levels in dB        Operation
       125                         Operators on a drill jumbo (Pneumatic)
       120                         Operator on a jack leg
       115                         Operators on a drill jumbo (Hydraulic)
       110                         1 m from a 40 kW axial flow fan
       105                         1 m away from a 20 kW axial flow fan
       100                         Typical heavy repair shop
       90                          Heavy trucks at 5 m
       80                          Busy office with tabulating machines
       70                          Loud radio in an average room
       60                          Typical restaurant
       50                          General office
       40                          Typical suburban area

                                        CHAPTER 4
       Mining is a huge industry worldwide. The term mining includes from operations
employing tens of thousands of people and moving millions of tons of ore and rock per
month, to an individual with a gold pan. Mining is carried out in almost all conceivable
locations, from tropical jungles to the high Artic, from 4 000 m above sea level to almost
4000 m below surface. A vast range of minerals is mined, requiring very different extraction
and processing operations. The impacted areas are generally identified as follows:
       1.    Energy consumption
       2.    Air
       3.    Water
       4.    Land, and
       5.    Health and safety.

4.1    Energy Consumption
       Mining industry is a large consumer of electricity. For example, in the USA in 1994,
the mining industry, including primary metal production, consumed 158 TWh (1012 Wh) of
electricity, which was 5% of total electricity consumption. In South Africa, in 1995, the
mining industry accounted for approximately 25% of electricity consumption. The reason
for these figures lies in the quantities of ore and rock that have to be transported by the
industry, resulting in huge vehicles or extensive hoisting systems for underground mines.
Cooling of deep underground mines is very energy intensive, as is pneumatic equipment,
which is used extensively. Smelting of many metals requires large amounts energy.

4.2    Air
        Surface mines may produce dust from blasting operations and haul roads. Many
coalmines release methane, which is a greenhouse gas. However, methane is generally
captured, where it is economically feasible to do so. Since some coal plants use CFCs,
HCFCs and HFCs, there is the potential for the release of these ozone-depleting substances,
but such releases are tiny. Tailing dams, if not vegetated or capped, may also be a source of
dust. Radiation is emitted from tailings dams where radioactive elements are found in the
ore. Smelter operations with insufficient safeguards in place have the potential to pollute the
air with heavy metals, sulfur dioxide, and other pollutants.

4.3    Water
               The mining industry uses large quantities of water, though some mines are
able to reuse much of their water intake. Mining brings sulfur-containing minerals into the
presence of air, where they are oxidized and react with water to form sulfuric acid. This,
together with various trace elements, impacts ground water, both from surface and
underground mines. Tailings dams and waste rock heaps are also sources of acidic drainage
water, affecting surface and underground water. The chemical deposits left behind by
explosives are usually toxic, and they contaminate and increase the salinity of mine water.
In-situ mining, in which a solvent is allowed to percolate through unmined rock, leaching
minerals directly, has the potential to contaminate ground water. Artisanal mining may
impact water where mercury is used to process gold.

4.4    Land
        Mining moves large quantities of rock. In surface mining, land impacts are very
large. These impacts may be temporary where the mining company returns the rock and
overburden to the pit from which they were extracted. Many copper mines, for example,
extract ore that contains less than 1 % copper. For many nonferrous metals, all of the mined
ore thus becomes waste. Artisanal mining, alluvial mining for gold and diamond often has
an impact far greater than the size of the operation. Many areas are marked by thousands of
small holes, which have been indiscriminately dug in search of precious minerals. Trenches
that scar the landscape are problems in some places. These activities may lead to erosion
and the localized destruction of riverbanks.

4.5     Health and Safety
       Mining operations range from extremely hazardous to being as safe or as dangerous
as any other large-scale industrial activity. Underground mining is generally more
hazardous than surface mining because of the poorer ventilation and visibility and the
danger of rock-falls. The greatest health risks arise from dust, which may lead to respiratory
problems, and from exposure to radiation (where applicable). The use of mercury for gold
extraction in artisanal mining is very hazardous to health and fertility.

4.6    Wastes
Wastes from the mineral resource based industries -- mining, milling, and metallurgy -- are
characterized as follows.

Mining Industry
Solid Wastes: The major solid waste disposal problem in the mining industry is the
handling and relocation of overburden and gangue. The rate of production of overburden
and gangue increases annually as the depth of ore bodies mined becomes greater and grade
becomes lower. Greater overburden problems are associated with open pit or strip mining
methods. The runoff from such solid waste piles is also of interest. Stream sedimentation
may or may not be a problem downstream from such disposal sites, depending on the nature
of the solid waste and on the nature of the climate in the mined area. The most common
methods of minimizing the effect of such solid waste disposal features on stream channels
are revegetative stabilization and landscaping.

Liquid Wastes: The major and most significant source of liquid waste in the mining
industry is acid mine drainage. Acid mine drainage is common in area where mine openings
intersect the water table and where the rocks contain pyrite or, less commonly, certain other
sulfides. Many coal seams have pyrite associated with them. Acid runoff from relocated
overburden is also a problem for the coal mining industry. The low pH water is produced by
the sulfides associated with the overburden, which oxidize upon exposure to the

Gaseous Wastes: With the exception of relatively insignificant production of dust, the
mining industry is blessed with few air pollution problems outside the mine. The major
gaseous problem confronting the mining industry is the production of methane, but this
problem is restricted to the atmosphere in underground openings, mostly in coalmines.

Milling Industry
Solid Wastes: Solid wastes constitute the major disposal problem of the milling industry.
The wastes consist of the host rock for the minerals that have been removed by various
concentration processes. The wastes are known as tailings and they are deposited in tailing
ponds. The solids are allowed to settle in the settling basin, and the effluent is decanted
from the cleanest portion. The solid residues pile themselves, if not properly handled,
constitute sources of dust and sediment to be distributed to streams by runoff.
Liquid Wastes: The major liquid waste disposal problem confronting the milling industry
consists of water decanted from the tailing ponds. The decanted water contains some
suspended solids and sometimes low concentrations of cyanide and other dissolved ions.

One challenging area of research for the milling industry and associated government and
academic institutions consists of the development reagents that are either nontoxic or at
least have minimum toxicity, particularly with respect to aquatic species.
Gaseous Wastes: With the exception of a few wet scrubbers and dust derived from the
tailings piles, the milling industry has negligible pollution problems of this type. Gaseous
wastes are essentially nonexistent in the milling industry.

Metallurgical Industry
Solid Wastes: The term metallurgical industry is broad and covers many industries. It is
difficult to define precisely. Moreover, many operations are connected with mines or mills
or both. The solid wastes from metallurgical processes range from innocuous slag from
blast furnaces to unstable solids produced by the washing of gaseous wastes and metallic
products. The most common are those that are the result of air pollution control devices
such as scrubbers and cyclones used for dust control. These solid wastes are usually in the
form of sludge from wet collecting devices, and they are frequently discarded in landfills.
Liquid Wastes: Liquid wastes from metallurgical industry range from wash water of the
raw material to effluent from wet scrubbers used for dust control.
Gaseous Wastes: Gaseous waste control is a continuing problem to the metallurgical
industry because all metallurgical processes produce gaseous emissions of one kind or
another. They may contain only particulate matter, they may contain only gases, or they
may contain both. The current practice is wet scrubbing to remove particulate matter,
followed by appropriate additional treatment to remove whatever gas happens to be present.
In the case of sulfur dioxide gases from copper and lead smelting operations, the gas
emitted is sulfur dioxide. The standard technique for handling sulfur dioxide is conversion
to sulfuric acid, which is now used in fertilizer production.

4.7    Pollutions
       In developing countries, which do not have a well-developed manufacturing sector,
the exploitation of mineral resources is of major importance from the standpoint of: (a)
contribution to the gross national product; (b) earnings of foreign exchange; and (c)
provision of job opportunities at all levels. The development of any industry, which results
in improvement of the national economy and the living standard of the people, should not
and need not restrict any segment of the population from enjoying a safe and pleasant

       Throughout the world, since the early days of mineral resource development, the
mining industry acquired the reputation of being a major and serious polluter of the
environment. Over the years, mining methods and equipment have been developed to
increase efficiency and to lower costs of metal recovery. They also lessen the extent of
pollution. However, the demand for metals and thus the exploration and exploitation of
mineral resources is continually increasing. Thus, although the pollution per ton of ore
mined has decreases, the continued increase in tonnages mined and processed can still result
in considerable degradation of the environment.
       In the development of environmental impact studies, needed to obtain permits for
exploration and for mine, mill and processing plant construction and operation, the
following stages in development from conception to close down of the proposed mining
enterprise should be considered:
       (a) Exploration
       (b) Pre-operational phase – site location and construction and operational process
       (c) Operational phase
       (d) Post-operational phase. Pollutions and degradation of the environment occur at
           each of these stages, and so will abatement measures be needed.
       From over-all standpoint and considering the different methods of mining and metal
processing and various metals extracted, a summary of the environmental hazards
associated with mining is identified below. They can be grouped into chemical pollutions,
physical pollutions and aesthetic impairment of the habitat.

4.7.1 Chemical Pollutions
Surface water pollution of public water bodies
       Water pollution of streams and lakes results mainly from liquid mine, mill or
processing plant effluents which may contain toxic elements, particularly metals such as
lead, zinc, cadmium, nickel, chromium, mercury and others. In some cases, the effluents
may contain fluorides, chlorides, nitrates and arsenic or cyanide plus organic processing
reagents such as frothers, collectors, flocculants, coagulants, dispersants and depressants.
Many of these elements are directly toxic to humans, animals and fish and to plant life, both
surface and aquatic. Other elements, in low concentrations, do not appear to be directly
toxic to fish and most plants. But they have serious toxic effects when concentrated in fish
and plants consumed by humans and animals.

       The tendency to acidity (low pH) of liquid effluents from many mine/ mill
operations can also present a serious environmental hazard both directly and indirectly
when such effluents are discharged without treatment into surface water bodies. The most
common direct result is the effect on aquatic life. Most fresh water fish require water in a
pH range between 5.0 and 8.5, usually the closer to neutral the better. Sudden variations of
the pH within this range can seriously affect their multiplication and growth. At a pH below
4.0 or above approximately 9.0, the fish die. An indirect but equally serious effect is that
acidic water reacts with heavy metals either in the effluent, in the soil or rock in the banks
or bottoms of lakes and rivers or underground. Then it renders the metals soluble with
resulting possible toxic pollution of both surface and underground water.

Underground water pollution
       This refers to pollution of the aquifers. Underground water can be contaminated by
the same pollutants as surface water due to mineral resource development. It s main cause is
the percolation downward through unconsolidated subsoil or faults in rock formations of:
(a) mine drainage water; (b) boreholes; and (c) drainage from mine waste disposal areas
(tailing ponds).
       If adequate controls for the discharge of liquid effluents are established and
implemented, active mines should not be significant sources of water pollution. However, in
many cases little attention has been given to this aspect of abandoned mines. If control
procedures are not implemented, pollution of the underground water by the causes noted
above can continue long after the mine/ mill site is abandoned. The United States
Environmental Protection Agency estimates that most significant water pollution problems
caused by mining in the United States result from abandoned mines of all types.

Chemical pollution of the atmosphere
       Air pollution is caused mainly by sulfur dioxide gases, nitrous oxide gases and
particulates mainly of lead, zinc, cadmium, arsenic, fluoride and silica resulting from the
processing of both ferrous and non ferrous metals by smelting.
       The environmental hazards caused by these air-borne pollutants are both direct and
indirect. Sulfur dioxide gas and nitrous oxides can cause respiratory problems in both
humans and animals at quite low concentrations. Further more, their effect on plant growth
is very serious. Thus the complete destruction of all plant life several miles around smelters
is common.

       SO2 emissions also cause accelerated corrosion of steel structures and deterioration
and discoloration of many natural stones, cement and paint. Particulates of lead, zinc,
cadmium and fluoride in most cases do not appear to have a significant effect on plant
growth. When plant material or water contaminated by these toxic metal particulates is
consumed by humans or animals, however, it can cause very serious problems.
       Probably the most widespread environmental degradation caused by chemical
pollution of the atmosphere associated with mineral resource development is the
phenomenon of acid rain. SO2 emitted into the atmosphere combine with water particles to
form sulfuric acid (H2 SO4). This acidified atmospheric moisture falls to the earth as acid
rain, often at considerable distances from the source of pollution. For example, the acid rain
problem in eastern Canada and the eastern United States, resulting from prevailing westerly
winds, derives in part from sulfur fossil fuel burning industries and thermal electric power
generating plants in the highly industrialized areas adjacent to Detroit and Chicago. In the
State of Pennsylvania, rainstorms have registered a pH of 2.8, almost as acidic as vinegar.
       The first indication of pollution by acid is the complete killing of fish in fresh water
lakes. It is estimated that to date (1982) approximately 1000 previously good sport fishing
lakes in eastern Canada and the United States are now devoid of fish due to acid rain. Acid
rain also progressively increases soil acidity, with resulting adverse effects on plant growth.
       Even an underground aquifer can be polluted by acid rain. A cause for alarm in this
regard is that when this acidified water is used in municipal water systems, it can dissolve
copper or lead used in plumbing systems and make the water toxic for human consumption.
       The removal of sulfur dioxides and nitrogen oxides from flue gases of smelters and
fossil fuel burning industrial and power generating plants is very difficult and costly.
Current technology in this field is nit adequate. Because of this and due to one if the more
important environmental hazards associated with mineral resource development.

4.7.2 Physical Pollution
Pollution of surface public water bodies
       Rivers, lakes and oceans may be polluted by solid in suspension resulting from
mining, mineral beneficiation and/ or processing. This is a visible type of pollution and so
in the past has caused the most complaints by the general public. In many countries, it is the
problem, which has received the most attention by environmental control agencies.
       The most common and direct form of such physical pollution of water is turbidity
and silting up of rivers and lakes, and, where water for irrigation is drawn off these bodies

of water, the silting up of these irrigation systems and even of the land being irrigated. In
many countries, the annual costs of removing mine tailings silt from irrigation canals and of
compensating farmer for damage to their crops from toxic silt deposits are very significant.
       Equally important, particularly from a long-term standpoint, is the effect on fish and
aquatic plant life, caused by high levels of solids in suspension – both in fresh water and in
seawater. Water turbidity results in decreased light penetration which seriously affects the
food chain in a marine ecosystem. Directly, it has a serious effect on fish and particularly
crustaceans by interfering with their gill function. It also causes an avoidance reaction in
many fish species. Larvae of all animals are seriously affected.
       This problem of physical water pollution (turbidity) is of major significance,
particularly in Indonesia and Thailand, due to extensive and continually increasing offshore
tin dredging operations. Ocean floor nodule mining in the ESCAP region could also give
rise to problems in this regard.
       Physical pollution of water by solids in suspension in most cases does not present a
serious direct hazard to public health. It is a prime indicator, however, of inadequate water
management and control in the mining cycle. Where there is turbidity, there is a distinct
possibility of chemical pollution by acids or toxic metals, the real hazardous pollutants.

Desertification of land
       This type of physical degradation of the land may result from the following types of
   (a) Land may be inundated with waste materials because of inadequate control of the
       deposition of tailings from mineral beneficiation plants or the failure of confinement
       structures to contain those tailings, particularly under tropical typhoon conditions.
       Examples include tin mine tailings in Indonesia, Malaysia and Thailand and copper
       concentrator tailings in the Philippines.
   (b) Dump areas or piles of waste rock or spoil can be created from both underground
       and surface mining and from overburden from surface mining operations.
   (c) Large open pits or trenches may result from surface mining operations.
   (d) Erosion will result from disturbance of the natural surface drainage patterns and
       from deforestation in all phases of site development. Deforestation occurs during
       exploration, during the pre-operation stage of construction of access roads, power
       and water supply lines and plant and staff facilities, while the mine is operating, and
       after it is abandoned.

Physical pollution of the air
        Air pollution may be caused by any of the following factors:
   1.   Dust resulting from wind erosion of barren unconsolidated tailings. This can be of
        serious concern in arid and semi-arid areas.
   2.   Dust resulting from blasting operations in surface mining and in dry grinding of
        materials. Examples are mining from coal, ore and cement.
   3.   Dust from haulage roads in all phases of the mining industry.
   4.   Dust from dry processing operations, such as for coal.

4.7.3 Aesthetic Impairment of the Habitat
        The beauty of the environment may be impaired by:
   1.    Barren and unsightly undeveloped waste disposal areas;
   2.    Deep pits;
   3.    Subsidence and collapse of abandoned underground workings;
   4.    Abandoned mine/ mill sites;
   5.    Use of abandoned mining areas for uncontrolled rubbish and garbage dumps;
   6.    Dry muddy watercourses devoid of fish and aquatic life;
   7.    Deforestation.

   Other types of degradation of the environment include:
           1. Objectionable odors from reagents in the tailings and from polluted
           2. Noise from blasting and from transport equipment, particularly in surface
        In conclusion, these are the basic environmental problems associated with mineral
resources development. The question is how they can be abated or hopefully eliminated so
that this valuable natural resource can be developed without undue restriction.
        Fortunately, technology has been developed which, if properly applied, can to a very
large extent eliminate or at least significantly diminish many of these environmental
problems associated with mining and metal processing – and in a practical and economical
way. Humans can have the best from two worlds. They can have cake and eat it too.

                                       CHAPTER 5

5.1    The Need for Research and Development
       Throughout the world in early days of mineral resource development, little or no
consideration was given to its effect on the environment nor was there any pressure by the
public. Also that time, the method of mining, ore beneficiation and processing used, by their
nature, were very prone to cause severe pollution. Over the years, mining methods and
equipment have been developed mainly to increase efficiency and to lower costs of metal
recovery. In most cases, they also lessen the extent of pollution. Yet the greater demand of
metals and the trend to work low-grade ores can still result in considerable degradation of
environment. Thus, constant vigilance and extensive research and development are and will
be required, and the result of research and development must be put into practice, if
pollution from mineral resource development is to be kept within acceptable levels.

5.2    Waste Generation from Mining
       The first step in any experimental or applied research should be to identify, fully
analyze the problems, and to list them in order of priority according to the seriousness of
their effect and the extent of occurrence. Every stage in the mining process has impacts on
the environment. The resource cycle, showing from exploration to disposal is shown in
Figure 5.1.
       In general, extraction and processing have the greatest impacts. Mining typically
involves construction of roads and buildings, stripping of surface vegetation and soil,
sinking of mine shafts or excavation of large open pits, and storage of rock waste, tailings,
and other discarded materials. Mining processes also release, nitrogen oxides and sulfur
oxides, vaporized metals, and volatile organic compounds, as well as dust blown from piles
of waste rock and tailings. Effluents that are contaminated with metals, acids, chemical
reagents, salts, or radioactive materials may pollute bodies of water. They can be grouped
into physical pollution, chemical pollution and aesthetic impairment of the habitat.

   Earth                                                                           Recycle

                                                                                   Metals,        Products (cars,
                    Exploration                           Refine/process           chemicals,     appliances
                                                                                   cement, etc.   structures, etc.)

                                                         Waste             Waste


                                  Figure 5.1        The Resource Cycle

       Regarding the impacts and mining stages, the major components of mining projects,
which are typical sources of impact and potential problems, are identified in Table 5.1.

       For easy reviewing and making different approaches, environmental parameters are
usually classified into four categories formulated by Battelle Pacific Northwest Laboratories
for the United States Army Corps of Engineers. The classification is:
                   1. Natural physical resources
                   2. Natural ecological resources
                   3. Human/ economic development resources
                   4. Quality-of-life values (aesthetic and cultural).
The full list of classification is a long one. Environmental parameters or problems that
would be found in mining industry may not be exactly the same as that list. Modern impact
assessment systems use these parameters or environmental resources/ values.

Table 5-1. Examples of environmental implications of mine project component
Mine Project Components               Potential Impact to be Controlled and Minimized

Underground mining                    Surface subsidence and changes in land use and
                                      drainage;       water    table    contamination;     polluted
                                      groundwater disposal.
Open cut mining                       Temporary or permanent changes in land use and
                                      drainage; aquifer disturbance and contaminated
                                      groundwater disposal on the surface; final voids and
                                      surface     spoil       emplacements;     visual     effects,
                                      ecological disturbance; air, water and noise pollution
Ore/coal preparation facilities and from operations.
operations,      and      materials
handling and storage                  Air, water, noise and visual pollution: land use
                                      changes and conflicts.
Solid and Liquid waste handling
and disposal (rejects and tailings)   Air, water, noise and visual pollution; land use and
                                      drainage changes and conflicts; ecological.
Roads, rail and service corridors
                                      Air, water, noise and visual pollution; land use and
                                      drainage        changes     and     conflicts;     ecological
Projects (as a whole)                 disturbance.
                                      Sociological,       economic      and   cultural     changes
                                      affecting residents, local communities and towns;
Construction phase                    change to local and regional land use, air and water
                                      Air, water, noise and visual pollution; land use
                                      changes; socio-economic implications.

        As an example of impact assessment, Table 5-2 shows the types of significant
impacts on environmental resources usually resulting from mining operations and the
effects on the mining operations due to environmental constraints, and a rough
approximation of the magnitude of these effects.

          Table 5.2        Environmental parameters for analysis of mining projects

      Environmental                            Ecological                                    Quality-of-life
                        Physical resources                          Human-use values
        Resource                                resource                                        values

                      Soil (Erosion/sedimentation)

                      Manufacturing industries
                      Ground water hydrology
                      Surface water hydrology

                      Ground Water Quality
                      Surface water quality


                      Land transportation
                      Dedicated area uses
                      Mineral processing
                      Terrestrial wildlife

                      Aquatic biology


                      Flood control
                      Water supply

                      Public health

                      Air quality




  Type of


  Impacts on
                      3 3 3 3 3 -        -   - 2 2 3 3 2 1 2 - 3 -           - 2 3 -   -   - 3 - -     - 3

  Impact on
                      3 - 3 - 3 3 - 3 -          -   -    -   - 2 - 1 2 -    -   - 1 - 1 3 2 - - 1 2

           Note: Numbers indicated magnitude of significant effects:
           (3) = Major, (2) = intermediate, and (1) = significant

5.3       Energy from Coal
          Mining industry is a large consumer of electricity and other resources. In South
Africa, it accounts for 25 percent total electricity consumption, 91 percent of which is
generated by the combustion of coal. Therefore it is possible to say that any technologies
that reduce electricity consumption indirectly reduce impacts on land, water and air.
          In 1993, more than 3.4 billion tonnes of coal was produced, half of which was used
to generate over 46 percent of the world’s electricity. Most of the rest was used in industry
for steel making and manufacturing cement. At current extraction rates coal will be
available for centuries, in contrast to decades for oil and gas. In the future therefore, coal
will continue to be the major fuel for electricity generation and thus essential source of heat
and power for industry and homes.

5.3.1 Environmental Impacts
          Although coal is a secure and abundant source of energy, its use presents several
environmental problems               as does the use of all fuels. The two main environmental
problems associated with fossil fuel combustion are the release of carbon dioxide, possibly
resulting in global warming, and the release of sulphur and nitrogen oxides, leading to acid

precipitation and photochemical smog. Coal burning also releases significant amounts of
trace elements into the environment. For some of these elements, including arsenic,
beryllium, lithium, selenium, thorium, the emissions resulting from coal burning outweigh
emissions from all other sources combined. There are also extensive environmental and
health-related problems associated with coal mining and processing. These include acid
mine drainage, subsidence, underground fires and explosions, and “black lung” disease.
       Yet coal is the most abundant of the conventional fossil fuels, so it may be worth
investing in new approaches to render coal more environmentally acceptable and more
efficient as an energy source. To that end, a variety of “clean coal” technologies is being

5.3.2 Clean Coal Technologies
       Clean coal technologies cover the full spectrum of modern processes developed over
the past 30 years to improve the utilization of coal, ranging from technologies commercially
used to those in the demonstration process.

Pre-combustion Cleaning
       Coal washing can remove as many impurities as possible from the fuel before
burning it so that it will emit fewer pollutants into the atmosphere (although the impurities
must still be dealt with in other forms). The cleaning is accomplished through a variety of
mechanical, chemical, and biochemical techniques. For example, sulphur-digesting micro-
organisms that thrive in sulphur-rich hot springs can be used to remove both organic and
inorganic sulphur. One Australian method can produce ultra clean coal (less than 1% ash)
involving alternate leaching of coarse coal by alkali and acid.

Coal Transport
       Coal transport, whether it be by conveyor, road, rail or ship, is viewed by the public
as a cumbersome exercise. The long-term trend in coal transport and distribution is
definitely towards pseudo-fluid, i.e. coal slurries and in some cases pneumatic suspensions.
There are currently common ways to create coal slurries depending on the type of end-users
such as coal-water mixture (CWM) and coal-oil mixture (COM). Pneumatic conveying is
attractive for coal transport over short distance as in cement, ash, wheat or plastic beads.

Combustion-Enhancing Technologies
       Another approach is to enhance the efficiency with which the coal burns, thereby
reducing the sulphur dioxide (SO2) and nitrous oxide (NOx) emissions. Some popular
methods are pulverized firing (PF), gasification and integrated gasification and combined
cycle system (IGCC), and fluidized bed combustion (FBC). In FBC, the combustion
chamber contains finely divided particles such as sand or ash. The combustion air entering
from below lifts these particles until they form a bubbling bed, which behaves like a boiling
fluid. Start-up burners heat the bed to some 800-950 °C. When the coal is introduced into
the bed, it ignites and burns. The amount of nitrous oxide produced in an FBC is much
lower than the amount produced in conventional coal plants When limestone is added to the
bed (limestone injection), most of the SO2 also is chemically captured before being released
into the environment. Fluidized bed technology results in cleaner, more efficient
combustion. It has the drawback of leaving behind a solid residue, but in some cases, the
residue may have commercial value.

Co processing and Conversion Technologies
       Many existing industrial technologies rely on a fuel liquid form. It would be very
costly to convert existing machinery or develop new technologies that could accept a solid
fuel. Therefore, attention has been paid to the possibility of converting coal into a form that
could be used by current industrial equipment. Liquid and gaseous fuels produced
synthetically from coal are referred to as synfuels. One method is co processing, whereby
finely ground coal is mixed with water or oil to form slurry. Another approach is conversion
of the coal into gaseous or liquid from, that is, coal gasification and liquefaction.

Post-combustion Cleaning
       Many power plants have reduced their emissions by installing “sc-rubbers” that
remove sulphur from flue gases, i.e. flue gas desulphurisation (FGD). Lime or limestone is
used as a sorbent to scrub SO2 from emissions. These processes can remove more than 90%
of SO2. Oxides of nitrogen can be limited by using “low NOx” burners. Up to 90% of
emissions can be reduced through a combination of methods, including selective catalytic
reactors (SCR). Bag filters or electrostatic precipitators, removing 99.9% of the particles,
and eliminating the black smoke, can collect dust, which leaves the system with the flue
gases. In electrostatic precipitator, the flue gases are passed between collecting plates where
dust is collected by means of an electric field.

5.3.3 Energy Alternatives
       The typical reservoir rock for petroleum hydrocarbons is a rock of high porosity and
permeability, such as a sandstone or limestone. However, other types of sedimentary rocks,
such as shale, coal seams, and low-permeable sandstone can also contain oil and gas in
appreciable quantities. The tar sands and oil shale are the most promising unconventional
hydrocarbon fuels that occur in an atypical reservoir such as a tight sandstone or
geopressurized zone.
       The largest known occurrence of tar sands is in Alberta, Canada. It covers 5,000 km2
and reaches a thickness of 60 m. It has been estimated to contain as much as 600 billion
                                    __                                        __
barrels of petroleum. The costs          both economic and environmental           of mining and
treating tar sands are high, but the process is technologically feasible. The oil sands could
provide all of Canada's oil needs for 475 years.
       Another potential source of petroleum is a waxlike organic substance called
kerogen, which forms when organic material is buried, compacted, and cemented in very-
fine-grained sedimentary rocks as marine and lacustrine (lake) shale. If burial temperatures
are not high enough to initiate the chemical breakdowns leading to the formation of oil and
natural gas, kerogen may be formed instead. If the kerosene is heated, it breaks down and
forms liquid and gaseous hydrocarbons similar to those found in oil and gas. All shale
contain some kerosene, but to be considered an energy resource the kerosene in an oil shale
must yield more energy than is required to mine and heat it. Because the energy needed to
mine and process a ton of shale is equivalent to that created by burning 40 liters of oil, only
shale that yield 40 or more liters of distillate per ton can be considered.
       Far more energy is available in the Earth’s energy than humans can use. What is not
clear is when humans will learn how to tap the different energy sources in ways that are
economical and do not disrupt the environment. Three sources of energy other than fossil
fuels have been developed to some extent in different parts of the world. These are biomass
energy, hydroelectric energy, and nuclear energy. Others, including solar energy, ocean
energy, wind energy, and geothermal energy have been tested and developed only on a
limited basis. Ocean energy encompassed tidal and wave energy as well as ocean thermal
energy conversion (OTEC).

5.4        Mining
       As much mining requires the excavation and removal of vast quantities of soil and
rock, this results in visual impacts on the environment, though these are of a short-term

nature. However, the action of disturbing geological strata brings rock into contact with air,
which initiates processes that result in environmental impacts. Some of the innovative
means that the mining industry is investigating in order to reduce these impacts are
presented in this section.

5.4.1 Non-explosive Rock Breaking
       Although blasting is very energy-efficient, the gases that the explosives generate and
the chemical deposits they leave behind are usually toxic, contaminate air underground, and
increase the salinity of existing mine water. A greater degree of control can be obtained
with mechanical systems rather with explosives in rock breaking. As a result, dilution is
reduced, and less energy is required to hoist and process the reduced quantities of rock, and
less waste material to be disposed of. The ability to work continuously, compared with the
delays required by drill and blast mining, brings major economic advantage to mechanical
mining. Mechanization can reduce worker exposure to dangerous conditions and therefore
improves safety levels.

Mechanical Cutters
       They have been used in the form of continuous and long wall mining machines by
the coal industry for several decades. They are now in use all over the world. Coal cutting
machines are a relatively mature technology. However, the braking of harder rock poses
problems on cutting tools. Another constraint in underground mining is the need to build
machines as small as possible. The fine quartzite particles in mines also lead to rapid wear
of machine parts, adding to the extent of the challenge.

Tunnel-boring Machines (TBMs)
       They were first introduced in the 1950s, when they were used to cut reasonably soft,
stable rock. Recent advances in cutter technology          in terms of diameter, load capacity,
gauge velocity, geometry, and physical metallurgy          mean that there are few rocks in the
world that may not be bored with TBMs.
       They can also be used in poor rock conditions. The chief drawbacks of TBMs, apart
from their high initial cost, are their inflexibility and their turning radii. However, where
their use is appropriate they bring the general environmental and safety benefits of

Raise-boring Machines
      They have also been in use for many years to cut vertical raises. More recently, they
have been used to cut short, horizontal tunnels as well. Their advantages over TBMs lie in
their lower electricity consumption (and therefore lower environmental impacts), lower
capital costs, less preparation work, shorter equipment set up times, reduced ground
support, and increased safety.

Steel Wire Ropes
      Steel wire ropes studded with industrial and synthetic diamond coated beads have
been used since the 1970s to quarry marble and granite. They are now being developed to
cut gold platinum stopes 20 - 60m thick. Slots are simultaneously cut above and below the
reef and the ore is blasted out with a light charge, such as detonating fuse. As diamond wire
mining requires a tunnel to be continuously developed ahead of the face being mined, such
that it always leads the face by about 20 metres, this method is not applicable to ultra-deep
mines. However, a circular diamond saw which does not require such development, thought
it does need an additional clearance slot to accommodate the saw unit, is under investigation
for possible use at depth.

Water-driven Hydraulic Hammer
       This impact mining system is currently at the production trial development stage. It
breaks and loads rock onto a reciprocating flight conveyor. Research is being directed to
increase the energy of the hammer from 4.5 to 8 kJ or more, developing a simpler and more
reliable hammer, and increasing the stoping width that can be mined. In addition to the
benefits of mechanization, the use of hydropower instead of compressed air eliminates the
use of electric power with the attendant environmental advantages.

Plasma Blasting
       The technology could be incorporated into a mobile mining machine capable of
drilling, blasting, and removing the fragmented rock on a continuous basis. The technology
uses the transformation of an electrolyte solution in a borehole by electrical energy into a
high pressure, high temperature plasma. The plasma produced by the transferred energy
expands faster than the normal seismic wave propagation velocity and forms a shock wave.
The shock wave in turn produces a stress field that shelters the rock without producing
excessive dust and debris.

        This energy transfer into a confined space is a process quite similar to a lightning
strike on a tree or rock. The peak power level is actually higher than that required by the
electrical network of a big city but the time span is so short that the electrical energy for a
blast costs only a fraction of a cent.

Laser Water Jet
        This is the controlled application of pulsed laser energy for hard rock excavation. It
does not involve rock drilling, and its main advantage over conventional techniques is the
ability to continuously excavate small particles from the rock surface. A cutter head fitted to
a jumbo would be supplied with laser energy via optic fibre cable from a remote laser unit,
and with a high-pressure water supply. The ore chips produced would be removed by
vacuum and pumped via slurry line to the process plant. The high-pressure water jet
provides a consistently clear path to the rock face through the dust, fog and other debris
normally found at mining face. The excavation mechanism involves vaporizing micron-
thick layers of materials from the surface in nanosecond (10-9 sec) time intervals. Provided
the expanding gases are held shortly against the rock surface with water, pressure
approaching 1,000,000 psi is generated. These pressures impart instantaneous shock wave
to the rock in much the same way as the traditional explosives.

The Penetrating Cone Fracture (PCF) Technique
        This technology uses a non-explosive chemical propellant to generate very high
transient gas pressures at the bottom of a short drill hole. The injector is a gun-like device in
which a propellant charge is burned in a combustion chamber external to the working face.
A propellant cartridge is inserted into a breech and fired using conventional technology. The
burnt propellant gases are directed to the bottom of a shallow drill hole by a barrel to create
a rapid but non-explosive pressure pulse at the hole bottom. It generates very low velocity
fly rock and size distribution of rock may be controlled. While much smaller quantities of
rock are broken per blast than with explosives, PCF may be used continuously since the
area does not have to be cleared in order to allow fumes to be purged.

Water Gun
        The great depths at which mining takes place (sometimes 4000 m below surface) can
be exploited in some innovative rock-breaking techniques that take advantage of the highly
fractured and stressed rock. It is hoped to exploit the water pressure at depth with a “water

gun”, which is at an early stage of development. Utilizing the principles similar to those
used by PCF technology, the barrel of the machine is inserted into a hole that has been filled
with water. Then very high transient pressures are applied, breaking the fractured rock. An
energy release of 35 kJ has been achieved in laboratory trial.

The Cardox Tube
        The tube consists of an alloy steel tube containing a charge of liquid CO2, with a
discharge head one end, and a firing head at the other end. At the firing end is a filling valve
through which the CO2 is loaded and low-tension electrical connections to detonate a
chemical energizer, which is also loaded into the tube. At the discharge end is carefully
designed steel rupture disc to contain the charge and external retaining pawls to hold the
tube in the rock. A hole is drilled in the material to be broken, and the loaded tube inserted.
The increased heaving mass of gas is discharged into the surrounding material which breaks
down along the least line of resistance with negligible amount of vibration and dust.

5.4.2          Hydraulic Drilling and Hoisting
        Hydropower rock drills have been used commercially since 1991. Hydropower
hydraulic rock drill systems, which require ancillary electricity only to pump used water to
the surface, use less energy. They have been shown to achieve better performance than
pneumatic drills; because the have better penetration rates than pneumatic drills at lower
operating costs. The used of water as a transport medium for broken rock is being
investigated as it has several potential benefits over the traditional hoisting method using a
wire-suspended skip. Hydraulic hoisting might be combined with underground crushing,
and possibly pre-processing of ore, thus introducing further efficiencies.

5.4.3          Ventilation and Cooling
        In South Africa, virgin rock temperatures of 60C are being experienced at less than
3000m below surface. It is necessary to cool air to a wet bulb temperature of no more than
28C to ensure safe, productive working conditions. Improved mine layouts are being
investigated to reduce the volumes of chilled air required. Although mechanization has
benefits, they generate heat and place load on ventilation system. Efforts have to be focused
on minimizing their heat loads. Chilled water has been used for cooling for many years, but
the use of ice is being introduced because it has better cooling properties. A new

desalination technique has made it possible to freeze underground mine water. Not only is
this a better use of a scarce resource, but it also reduces the environmental impacts and costs
associated with purifying the water before use.
        New machines use either non-CFC refrigerants or water itself as refrigerant.
Ammonia, which does not deplete the ozone layer, is also used in some water-chilling and
ice-making plants above ground. However, because of its toxicity, corrosivity and
flammability at certain concentrations, it is not used in plants installed underground.
        The insulation of intake airways is also being investigated as a means of reducing
the loos of energy to the rock forming the airways. This will reduce the electricity
requirements of the ventilation plant.

5.4.4          Support
        Approximately 1.2 M tons of timber is used annually underground in South African
mines to support the roof. It is a good practice to remove this combustible product from
underground mines. Backfilling has several environmental, safety and economic benefits. It
reduces the surface waste, ventilation and cooling requirements. Hydraulic props are able to
withstand blasting conditions, and may be installed before blasting. Rock bolts and other
support combination also have environmental benefits, compared with timber.

5.5        Processing
5.5.1          In situ Extraction
        In concept, the underground ore body is first fractured with explosives, then the
laxiviant (leaching) solution is allowed to percolate through the ore, being introduced from
the surface by way of injection wells. The pregnant solution is allowed to collect in a drift
and is pumped to surface, where it is processed. By careful design and good underground
management, and under favourable geological conditions, the laxiviant may be kept
separate from the ambient groundwater. This method offers the advantages of minimal
surface disruption and solid waste generation, and low operating and capital costs.

5.5.2          Biological Oxidation
        The use of Thiobacillus bacteria for the extraction of a variety of metals from their
ores is currently at various stages of development. They have been used in the processing of
refractory gold ores for a decade, have recently been applied to nickel, and are being
investigated in connection with copper ores. The bacteria consume sulphide in pyritic ores

by oxidizing it, thereby making some metals accessible. Otherwise, they would require
expensive and polluting extractive processes. The traditional methods of processing pyritic
minerals require roasting the ore, which may produce arsenic and sulphur dioxide. It is
oxidized and reacts with water to form acid rain. Pressure oxidization, which does not
produce these pollutants, is much more expensive than biological oxidative process.

5.5.3           Reduced Consumption and Improved Recycling of Cyanide
        Cyanide is used in the processing of many gold ores. While it is effective, it is also
very toxic to humans and animals. Cyanide is toxic when inhaled, absorbed through the
skin, or swallowed, like arsenic, which may remain in the mine spoils after the leaching
process has been completed. The main challenge in cyanide leaching is to prevent negative
environmental and health impacts. Workers must wear eye protection and impenetrable
clothing when dealing with cyanide, and work area must be well ventilated. Any cyanide
and cyanide-metal complexes must be removed from effluents before they are released in to
the environment. Several treatment technologies that are over 95 percent effective in
removing cyanide from solution are available.
        Research is being undertaken to minimize the consumption and improve the
recycling of cyanide by means of different recovery methods. While most cyanide is
believed to be destroyed by sunlight on tailing dams, it makes environmental sense to
reduce the amounts of cyanide in the processing of ores. Now that Myanmar has granted a
lot of artisanal gold mining, environmental and health impacts are at dangerous level, and
deserve effective research and solution.

5.5.4           Regeneration of Activated Carbon
        The carbon-in-pulp process was introduced in the late 1970s and resulted in marked
increases in extraction efficiency. The activated carbon used in the process is regenerated
using a rotary kiln or direct resistive heating. Research is currently under way to see if
regeneration is viable using microwave technology instead. Based on laboratory-scale
experiments, indications are that this method will have lower operating cost (primarily
because of its reduced energy consumption) but higher capital cost than the established

5.6           Waste Treatment and Disposal
         It is said that waste is resource out of place. Waste consists of the residual materials
and sources and by products that are generated by human use of earth resources. According
to this definition, only some kinds of wastes are pollutants or contaminants         materials that
have harmful impacts and degrade the environment. Other kinds of wastes are not
pollutants; they are merely leftovers that have not yet found their way to an appropriate
         Mining generates a very large volume of waste, virtually all of which is disposed at
the mining or processing site. Some of the solid waste is in the form of piles of discarded
rock, or gangue. Another major component of mining wastes is tailings, the slag and sludge
that are left behind after processing or smelting. Piles of waste rock and tailings are an
eyesore, and the minerals in them may combine with rainwater to from acidic runoff. Dust
can be blown from the piles and create problems for neighbouring communities. Waste
management options can be:
         1.     Waste avoidance
         2.     Waste reduction
         3.     Waste reuse (direct reuse of waste materials)
         4.     Waste recycling (using valuable components in other process)
         5.     Waste treatment, and
         6. Waste disposal
                The technologies and processes described above have been aimed, directly or
indirectly, at options 1 to 4. Some methods focusing on options 5 and 6 are described

5.6.1            Semi-dry Tailings Disposal
         A large proportion of mining waste consists of tailings. They are usually fine
particles, and are commonly deposited as slurry to form huge tailings dams or mine dumps.
Semi-dry stacking of tailings was introduced in the 1980s. It is likely to grow in popularity,
as it offers the following advantages over conventional, wet disposal:
         1.     The density of the deposit is much higher, with the result that it occupies a
                smaller volume and surface area;
         2.     Shear strength is considerably increased, and thus the deposit is much more
                stable and safer, and reclamation is easier;

        3.   Reclamation and revegetation activities may be carried out almost immediately
             after deposition has finished;
        4.   The dam may safely be built to a height which would not otherwise be
        5.   The risk of groundwater contamination is significantly reduced.

5.6.2          Co-disposal of Coarse and Fine Residues
        Another tailing disposal method that offers environmental benefits is the co-disposal
of tailings with waste rock, rather than the traditional method of disposing them separately.
The method necessitates a very low water content of the tailings. It has the following
advantages over wet disposal:
        1.         Water consumption is reduced;
        2.         The lack of free water means the potential for groundwater pollution is
        3.         Provided the foundation strata are adequately strong, there are no
                   potential problems of slope stability;
        4.         Erosion losses caused by both wind and water are reduced; and
        5.         The permeability of air and water is reduced. Thus, the likelihood of
                   spontaneous combustion and the rate of mine drainage are reduced in
                   deposits where these problems occur.

5.6.3          Control of Dust Emissions from Tailings Dams
        Techniques for vegetating tailing dams have been developed over many years, in
order to improve their appearance and control dust. Because mine tailings include fine
particles, there is great potential for dust pollution from tailing dams before they are
reclaimed or revegetated. In order to inhibit dust emissions from tailing dams, research is
being made to understand the mechanisms that control the generation of dust by wind and
other means. By being able to model dust fields, it is hoped that it will be possible to
improve the design of tailing dams.

5.6.4          Waste Disposal in Underground Old Workings
        Considerable interest was being shown in the underground disposal of waste in
Germany, particularly as existing landfill sites were approaching the end of their lives and
the establishment of new sites was meeting strong opposition from local population. Mining

industry can help in this situation. Raw domestic refuse is unacceptable underground.
However, incineration yields residues, which can be readily stowed in abandoned workings.
HCL is removed by injection of ground-limestone, slaked lime, or caustic soda. The residue
from the process is mixed with fly ash before disposal.
       Underground disposal is an ideal way of dealing with the mixed waste since the
danger of soluble chlorides leaching into the environment is eliminated. In Germany,
licenses have been obtained for two materials:
   1. Fly ash mixed with flue-gas-cleaning residues from incinerators, and
   2. Insoluble matter from the re-saturation of brine for electrolysis using rock salt

5.6.5 Underground Sludge Stowing
       To evade the construction of a tailing dam and integrated problems, underground
sediment recharge method can be used. This method can be applied to any abandoned mine
having goafs, galleries, or other caverns below the level of the adit out of which mine water
is flowing. As shown in Figure 5.1, even a cavern lying above the outflow level at present
may be used as a sludge dumping area if the outflow level can be raised above it by closure
of the adit. Additionally, a decrease in quantity of mine water and improvement of water
quality may be expected.

                                        Cave -in

                                                             Outflow level raised by closure of
                                                             the adit (X)


                                                   Cavern to be stowed

                                                   Cavern made rechargable by closure of the adit

                                              x    Closed

                     Figure 5.1. Cavities to be recharged with sediment

        The problems of the method are the settling characteristics of the sludge, the
possibility of re-dissolution, and the risk of blowout; however, these were almost resolved
according to the research conducted by the Agent from 1983 to1988.

5.6.6          Commercial Use of Coal Seam Methane Gas
        As recently as 20 years ago, the coal industry generally regarded methane gas as a
hazardous nuisance that had to be controlled and removed to make underground mining
safer. Its commercial potential was ignored. In the late 1970s, federal and technological
advances stimulated seam gas exploration and development. A number of coal mines in the
United States have, in recent years, collected methane from seam drainage operations and
sold electricity produced by electricity.
        In Alabama, up to 40% of the state’s natural gas supply comes from coal seam
methane recovery wells. Endowed with vast amounts of coal resources, Australia is the next
hot spot for coal seam methane development. Small research and development projects are
scattered among China, Poland, and Russia, but progress has been slow.

5.6.7          Treatment of Acid Mine Drainage (AMD)
        When piles of waste rock or tailings are left exposed, the waste will eventually
interact with rainwater, groundwater, or surface water. If the waste pile contains coal or
sulphide minerals, this interaction is likely to produce sulphuric acid (HSO). The acidified
runoff from the waste pile is called acid mine drainage or acid runoff. Acid mine drainage
was first officially noted in 1698 in Pennsylvania coal mines. In addition to acids, mine
runoff can contain toxic contaminants such as heavy metals or cyanide. Bedrock geology
and soil chemistry are the main factors that determine whether a surface water body will be
vulnerable to acidification. A lake situated in granitic bedrock is more susceptible to
acidification than a lake situated in limestone. Calcium carbonate in the limestone helps
neutralize or buffer the effects of acids added to the lake water.
        In West Virginia, the passive treatment system, SAPS              successive alkalinity
producing system        was used by Anker Energy Co., to add alkalinity to an acid seep that
came from the backfill of a reclaimed area. First, the two seeps were captured and directed

towards the SAPS. From the surface, the SAPS looks like an ordinary pond. However,
under the water lies a system of compost, limestone, and piping designed to lower the
water’s acidity by adding alkalinity. The SAPS consists of a 4-ft bed of limestone, totalling
3 000 tons. Layered in the limestone are rows of perforated pipes that spread out equally
throughout the pond. A layer of compost lies on top of the limestone to remove the oxygen
from the water so the iron does not drop out of solution. If the iron drops out of solution, it
leaves a film that coats everything, which can clog the SAPS drains. This also impairs the
dissolution of the limestone.
       It is necessary to determine how many pounds of acidity must be neutralized and
design life. At the end of the design life, the limestone should dissolve, and new limestone
must be put to start again. Because the SAPS is a passive system, it is less labour intensive
than active treatment. Only periodic monitoring is required to assure that the system is
functioning properly.
       Other systems that Anker used were storm water diversion, improved active system,
and layers of CFB ash and refuse. The typical treatment chemicals used in active treatment
are: anhydrous ammonia, sodium hydroxide (caustic soda), hydrated lime, and pebble. CFB
ash, coal combustion by-products, is highly alkaline and tends to harden like cement. It has
a very low permeability Alternated layers of CFB ash and refuse can stop an extremely
acidic seepage that discharged into the Cheat River in Albright, West Virginia. Another
waste material used for reclamation by Anker was slag. At Osage surface mine, the
company filled ditches with 900 tons of slag from old steel mill to generate alkalinity. It is
an economical source of alkalinity and generates much more alkalinity than limestone.

5.6.8 Neutralization Methods
       After performing a comprehensive investigation of the properties of the mine water
(quantity, ingredients, and concentration), and other conditions, a mine water treatment
system is selected. Each mine employs a method unique to it. At present, the following
method is widely used.
       1. Adding slaked lime to acid mine water, to neutralize the water as well as to
           precipitate the heavy metal ions (Zn, Cu, Fe, etc.) dissolved in the water as
       2. Introducing the slurry, produced by neutralization reaction to a settling basin or
           thickener to perform solid-liquid separation, discharging the supernatant liquid

           into a river after its quality satisfies the effluent standard, and condensing the
       3. Pumping the thickened slurry to a dumping area without either further
           processing or transporting after dewatering by a filter press.
       As a means to reduce the quantity of sludge disposed to the dumping area, closing of
the level and return of the neutralized sludge to the mine can be done.
       Flow charts of the mine water treatment process of various neutralization methods
are shown in Figure 5.2.
(1) Conventional neutralization method

  Mine W ater

                        Neutralization          Solid-liquid
 Slaked Lim e                                                            Treated W ater
                        Reaction                Seperation

                                                     Sludge                Disposal

(2) Sludge return neutralization m ethod

   Mine W ater

                         Neutralization          Solid-liquid
  Slaked Lim e                                                           Treated W ater
                         Reaction                Seperation

                                                     Sludge                    Disposal

 (3) Sludge return reverse neutralization m ethod

   Mine W ater

                      Neutralization       Neutralization       Solid-liquid
  Slaked Lim e                                                                        Treated W ater
                      Reaction             Reaction             Seperation

                                                                Sludge                    Disposal

(4) High Density Sludge (HDS) M ethod

    Mine W ater

                       Conditioning          Neutralization       Soild-liquid
   Slaked Lim e                                                                           Treated W ater
                       Tank                  Reaction             Seperation

                                                                    Sludge                  Disposal

          Figure 5.2. Neutralization of mine water by slaked lime

5.7       Recycling
       As a means of conserving mineral resources, it is wise to substitute abundant for less
abundant resources. Another way is to recycle the existing metal products as a partial
alternative to mining increasingly lower grades of ore. Recycling of some metals may be
costly, particularly in collecting the dispersed scrap. However, the energy requirements for
processing may be proportionally much less. For example, recycling scrap aluminium
requires less than 5 percent of the energy used in producing primary molten metal from the
ore. In Australia, about one fifth of the total consumption is metal recovered from collected
scrap. If the metal recovered from scrap produced in smelting and processing is included,
then the figure would be 50 percent (Hore-lacy, 1976). In 1996, the reclaimed scrap of non-
fuel materials in the United States reached $14 B (Lowrie, 1997). During the period, 1975-
1980, the United Kingdom figures for copper were 211,000 tons/yr (33% of total
consumption), for zinc 70,000 tons/yr (22% of total consumption) and for iron 16 million
tons/yr (68% of total consumption).

                                        CHAPTER 6
       Mining can affect and sometimes destroy other land resources and the use of land
for other purposes. Mine rehabilitation has become the modern approach to mineral
resources management. A detail proposal for the rehabilitation of the site should be
contained in the mine plan.

6.1    Rehabilitation Objective
       Mining is only a temporary land used and a clear rehabilitation objective consistent
with the projected future land use of the area must be defined. From the beginning, this
objective should be established in consultation with relevant government departments, local
councils, landowners, etc.
       The long-term rehabilitation objective may vary considerably at different sites. In all
cases, the first objective will be to protect the safety and health of people living in areas
surrounding the site. The rehabilitation programme may involve:
       1. Restoration of the area so that the pre-mining conditions are replicated
       2. Reclamation of the area so that the pre-mining land use and ecological values
           canbe re-established in similar conditions
       3. Remodelling of the area so that it is returned to a use substantially different to
           that which existed before mining.
       The final objective of rehabilitation may be reclamation of the site to a safe, stable
and non-erodible condition. Whatever the final rehabilitation objective, rehabilitation plans
should be drawn up as early as possible and be an integral part of the mining plan.

6.2    Description of Site
       A survey of the site is essential to provide a baseline standard for later rehabilitation.
The significance of various factors will vary from site to site. A site survey will normally
need to include information on:

       1. Land form
       2. Geology
       3. Soil types
       4. Surface and ground water
       5. Flora and fauna components

         6. Land use, and
         7. Heritage or other specific conservation values.

6.3      Site Plan
         Many of the adverse impacts of mining operations can be avoided or reduced by
careful siting of the proposed operation. Preparation of a site plan enables the clear
identification of the important physical elements of the site including environmentally
sensitive locations. Optimum positioning of the site infrastructure and mining procedures
can then be decided upon to:
      1. Protect access to the ore body and any possible extensions
      2. Optimise haul distances for ore and overburden.
      3. Avoid the impact on environmentally sensitive locations
      4. Minimize disturbance beyond the mine excavation and avoid steep cuts and fills or
         other extensive earthworks
      5. Minimize the noise and visual impacts (including lights) on adjacent land users
      6. Develop a site drainage plan, and
      7. Optimise the configuration of the mined areas for proposed post mining land use. In
         areas where the potential for visual and noise impacts is large, provision to impacts
         and remedial works can be incorporated into the site plan. A carefully sited access
         point can reduce visual impact.

         Similarly, careful planning and design on pit development can reduce visual impact
as shown in Figure 6.1.

                       DIRECTION OF WORKING            DIRECTION OF WORKING

                          CRITICAL VIEW POINT                  CRITICAL VIEW POINT

                VISIBLE                                INVISIBLE

                             Figure 6.1. Planning pit development

       Another way is to use overburden or waste materials as barriers as in Figure 6.2.
Overburden can be shaped into mounds and vegetated to eliminate visual impacts.
Stockpiles can be concealed in the pit.

               Figure 6.2 Placement of waste material as barriers
       Landscaping of the entrance to mining or quarrying operations and progressive
rehabilitation (during mining) can be incorporated into these plans.

6.4    Making Safe
       Each mine will have particular characteristics that will influence the procedures
adopted in the rehabilitation programme. These characteristics may be obvious but critical
differences are often only identified by careful investigation. The proposed post mining land
use will also influence the procedures and plant species used for rehabilitation. After
planning, the first step in rehabilitation is to clean up and make the area to be rehabilitation
safe. This involves:
       1. Removal of infrastructure and unused or unwanted equipment. No facilities or
           equipment should remain on site unless with the written approval of the
           landowner or relevant authority.
       2. Removal of rubbish for disposal at approved sites. Particular care is required
           with residual toxic or hazardous materials including contaminated packaging and
       3. Removal of all services.
       4. Removal or burial of concrete slabs, footings, etc.
       5. Backfilling or securely and permanently covering any shafts, pits or similar
       6. Restricting or preventing public access by removal or closure of access roads
           and tracks.
       Because of rehabilitation, there may be the potential to create safety hazards. In
particular, revegetation may create a fire hazard. It is necessary to observe statutory

regulations and contact with local fire control authorities. A fire management plan for the
site should be prepared.

6.5      Landscaping
         The reshaping and grading of a site is an essential aspect of rehabilitation. Unless
slopes are stable, the effectiveness of subsequent topsoiling and revegetation is greatly
reduced and maintenance prolonged. The final landform should be hydrologically and if
possible, visually compatible with the surrounding area. When planning the final landform
the following factors should be considered.
      1. The stability of land form. The erosion potential of the material on site needs be
         assessed. A geotechnical engineer’s report may be required. Steep and long slopes
         allow surface runoff to accelerate resulting in erosion. Natural slopes existing in the
         area should be considered.
      2. The drainage pattern for the overall site must be planned as part of the overall
         landscaping. Drainage density of adjacent land areas and the geology of the site will
         need to be considered.
      3. Slope design. Slope should be designed as shown in Figure 6.3 to reduce the
         velocity of runoff as the catchment of the slope increases.

                 20 - 30%                   70 - 80%

                                                                              AVERAGE SLOPE ANGLE
                                        AVERAGE SLOPE                         TO SUIT MATERIAL &
                                        ANGLE TO SUIT                         LAND USE BUT LESS
                                        MATERIAL & LAND                       THAN 20 DEG.
                                        USE BUT LESS THAN
                                        20 DEG.

                                                                         BACK SLOPED BENCH
                                                                         (ABOUT 4 METERS
                                                                         WIDE) CONSTRUCTED
                                                                         WITH 3 DEG. SLOPE OR
                                                                         LESS ALONG CONTOUR

                 CONVEX                   CONCAVE

                            (a) Ideal slope profile              (b) Profile design when space limited

                                              Figure 6.3 stable slope forms

       Where site limitations prevent the formation of a stable slope profile, benches or
similar erosion control methods may be required. Slopes with an overall convex profile
should always be avoided. Benches are best located in the middle of the slope. Where long
slopes cannot be avoided, several benches may be required and their spacing will need to
consider slope and runoff characteristics.Topsoil will generally not adhere to slopes steeper
than 27. The maximum slope for mechanically spreading topsoil is approximately 19. The
maximum slopes considered suitable for different land uses are:
                   Forestry                              38
                   Hill grazing                          28
                   Improved pasture                      15
                   Some buildings and roads              12
                   Rotation cropping                     5
                   Housing                               3
       Depending on geology, soils and other site-specific variables, lesser slopes may be

6.6    Topsoil Management
       Topsoil is almost an essential factor in successful rehabilitation programmes
particularly during the period of initial plant growth. Subsoil conditions become of more
importance in the longer term. Topsoil (or weathered surface material) generally contains
seeds, nutrients and microorganisms that are essential to plant growth and if they are lost
then the system will generally take a longer time to re-establish. During the planting stages
for rehabilitation the following should be considered.
       1. The amount of topsoil. The top 100mm-300mm of soil should be recovered.
       2. Direct replacement of topsoil will give the best results because it prevents or
             reduces the deterioration of the biological components in the soil during storage.
       3. If stockpiling of topsoil cannot be avoided:
             (a)      Plan to re-use the topsoil as soon as possible.
             (b)      Do not store in large heaps. Low mounds no more than 1 to 2m high are
                      recommended. Revegetate the stockpile to protect the soil from erosion,
                      discourage weeds and maintain active populations of beneficial soil
             (c)      Locate the stockpiles where they will not be disturbed by future mining

           (d)    Consider how the topsoil will be re-spread.
       4. Soils should not be stripped when they are wet as this can lead to compaction
           and loss of structure
       5. The appropriate depth of topsoil will depend on the site. Approximately 0.2-
           0.3m of replacement soil desirable where overburden material is not toxic to
           plant growth. If adverse conditions exist, topsoil may need to be increased and
           other measures to isolate toxic material considered.
       6. Where there are only limited supplies of topsoil, priority areas must be
           identified. These are likely to be those areas more prone to erosion (lower end of
           slopes, embankments, etc) or locations where the physical and/or chemical
           characteristics of the overburden are adverse to plant growth (e.g. saline alkaline
           or acidic spoils, etc). These areas must be revegetated in strips.
       7. Augment topsoil with other material. An underlay of subsoil with reasonable
           properties for plant growth will produce better results than a thin layer (0.1m) of
           topsoil alone. Nevertheless, placing subsoil should be avoided near the surface,
           which is excessively sandy, cleyey, or which has pH below 5.0 or above 8.5,
           chloride 3.0% or electrical conductivity 400 millisiemens per metre.
       8. Assistance and advice should be sought on taking soil samples and the selection
           of a suitable laboratory.
       9. Care is required, when topsoil is taken from other mine sites, to ensure weed
           species are not introduced.
       For reference, the practices of soil improvement for vegetation conducted in Japan
are shown in Table 6.1.

                   Table 6.1. Improvement for sediment and soil covering

      Mine      Improvement for covering soil and heaped substance

                After applying covering soil on the                                 soil
                sediment, grass seeds were sown.
                A ditch with a depth of 30 cm was
                                                         .. .              .. ...
                                                           . ..           ..
   Taisho       dug, in which covering soil was               ..          .
  Nishimata     put, and then grass seeds were                  ...           Covering
                sown.                                            ...
                                                                  ...         soil

                Calcium hydroxide was sprinkled                                     Covering
                on slime at a rate of 0.8 to 1 kg/m 2,                                   soil
                covering soil was applied, and then                                  Calcium
                grass seeds were sown.                                              hydroxide

                The surface of slime was plowed to                                    Covering soil
                a depth of 20 cm, covering soil was                                    Plowed slime
      Myoho                                                                            layer, 20cm
                applied, and then grass seeds were
                                                                                    Slime dam
                sown.                                                               as such

6.7     Erosion Control
        Control of erosion is important during mining and rehabilitation. The effects of
erosion may require remedial works on sites. Inadequate control of erosion can lead to a
reduction in water quality downstream. A major objective of most rehabilitation
programmes is to establish an adequate cover of vegetation to stabilise the site.

Wind Erosion
        While a vegetation cover is being established there are three basic methods of
controlling wind erosion on disturbed soils.
        Protection of the soil surface by natural or manufactured materials or mulch. In most
cases, the use of these materials may form an integral part of the vegetation programme the
aim of which is to establish a permanent protective cover. Selection of the mulch will need

to consider availability of material, proposed method of seeding or planting, future
trafficking of the area, and effects of mulch colour on soil.
        Maintenance of the soil surface in an erosion resistant condition This means leaving
the surface lumpy. This is particularly difficult in sandy soils because aggregation of soil
particles is not strong to resist abrasion. Keeping the soil surface damp using sprays or
water tankers will increase the aggregation of particles and their resistance to wind erosion.
                In arid areas with sandy soil, creating a dimpled effect similar to the surface
of a golf ball has been shown to assist natural revegetation. The dimples are created by
using trucks rather than scrapers. The method is best suited to the tops of waste dumps or on
slopes of 12 or less.
        Reduction of wind velocity across the disturbed areas by establishing windbreaks
Windbreaks may be rows of trees or shrubs retained or planted at right angles to the
direction of the erosive winds. Trees or shrubs should be fast growing and hardy but be
cautious not to introduce weeds, etc. Artificial windbreaks, in the form of various kinds of
fencing, are also available. Selection and placement of windbreaks should consider the
direction of critical winds, height and spacing, permeability, and length and continuity. On
level ground, protection will extend a distance approximately 20 times the height of the

Water Erosion
        Erosion by water is caused mostly by surface runoff from intense rainfall events.
The important factors influencing runoff characteristics are rainfall, area of disturbance,
catchment area, slope and profiles of channels (angle, length, and cross section, etc.), soil
characteristics and land use.
For controlling water erosion, there are four basic methods as follow.
Minimising Area of Disturbance
        Clearing of vegetation should be limited to that is necessary for the safe operation of
the mine (including fire management). Minimising the area cleared will reduce costs for
both clearing and site rehabilitation. Provisions that will assist in minimising the area
cleared include:
        (a) Preparing detailed site and mine development plans
        (b) Restricting progressive clearing ahead of the pit to that required for no more than
            6 month's production

       (c) Clearly identifying on the ground, areas designated for clearing
       (d) Training mobile equipment operators on the need to identify the exact limit of
           area to be cleared prior to commencement
       (e) Close supervision of mobile equipment during clearing
       (f) Avoiding the use of heavy equipment to clear vegetation where topographic
           conditions and vegetation cover do not warrant it, and
       (g) Including penalty clauses for employees and contractors for damage to areas not
           designated for clearing.

Restricting Entry of Runoff to the Site
       Construction of diversion channels or holding structures such as banks, drains or
dams will effectively limit the entry of water to the site. This will reduce the potential for
soil erosion on the site. When planning diversion structures, the following should be
       (a) All structures, permanent or temporary, will need to be designed to
           accommodate anticipated peak flows.
       (b) Information needed to properly design major erosion control structures will
           include rainfall frequency and duration curves, catchment size, and runoff
       (c) A storm return period will need to be nominated as part of design criteria.
           Design parameters will depend on the purpose of the structure and its anticipated
       (d) For water storage and limiting entry of water to site, the dam must be adequately
           sized, and provision be made for safe discharge.
       (e) Waterways to divert runoff or accept flows from the site should be designed to
           avoid erosion within the channel. Longitudinal profiles should be similar to
           slopes and convex profiles should be avoided.
       (f) Triangular or trapezoid cross-sections are preferred and can be constructed with
           a small dozer or grader. Avoid rectangular or similar shaped cross-sections,
           which occur with backhoes, etc.
       (g) Dimensions of diversion channels need to ensure flow velocities, which will not
           damage the channel. In easily eroded soil, maximum safe velocity is
           approximately 0.5 m/sec increasing to 1.0 m/sec for erosion resistant soils.

           Where the channel is grassed or covered by other vegetation these velocities can
           be doubled.
       (h) Where flow velocities cannot be reduced to a safe level, it will be necessary to
           protect the channel with erosion resistant lining materials. Suitable materials
           include concrete, galvanised iron, plastics or rubber sheeting (old conveyor belts,
           etc.), jute and bitumen matting and timber.

Encouraging Infiltration
       This is most effectively achieved by ripping the disturbed area along the contours. In
addition to increasing infiltration, ripping relieves soil compaction, keys topsoil to subsoil,
and increases the volume of soil readily available to plant roots and provides places for
seeds to lodge. Important considerations when ripping include:
       (a) Always rip precisely along the contour. This will normally require a surveyed
       (b) Rip the area following re-contouring and topsoiling.
       (c) Ripping should be as deep as possible 0.8-1.8m depending on the material,
           available equipment and subsurface condition.
       (d) Do not rip when soil condition is too wet to allow the soil to shatter.
       (e) If ripping results in bringing large amount of rock to the surface, reduce of tines
           or discontinue.

Managing Water Leaving the Site
       Diversion of established drainage lines or watercourses would normally require
specific approval. Similarly diversion or discharge of runoff may have legal applications
and necessary approval or permits, etc. should be obtained. Other important considerations
       (a) Sediment dams are most commonly used to control and retain major and longer
           lasting sources of sediment-loaded run-off.
       (b) Most sediment will be carried by the infrequent high intensity rainfall events.
           Dams and spillways must be designed to meet the majority of those events or
           they will not function at the most critical times. Seek advice from appropriate

       (c) Ensure construction material is suitable. Leakage from the dam may in itself not
           be critical but consequent instability of the dam wall through percolation may
           result in failure.
       (d) Locate dams so that run-off is readily diverted and channel gradients are not
           excessive, and the need for protective measures within the channel is minimized.
       (e) In many dams sediment levels greater than half the depth will begin to
           significantly reduce overall capacity and should be removed.
       (f) Design of spillways, outlet pipes, etc. is an integral part of the dam design and
           function. Clarified water will need to be discharged from the settlement dam in
           readiness for subsequent storms. Figure 6.4 (Australian Mining Industry
           Council, Undated) illustrates some options.
       (g) Always ensure flow rates at the discharge points are reduced to safe levels. If
           necessary, install means for dissipating energy, e.g. suitable body of water,
           concrete structure rocks, etc.

                           LEVEL OF EMERGENCY          PIPE WILL COMMENCE
                           EXTERNAL SPILLWAY           SIPHONING WHEN WATER
                                                       REACHES THIS LEVEL

                                                          SEDIMENT BUILD UP

                           GATE VALVE
                                                                  INLET RAISER & SCREEN
               A. Self-siphoning pipe


                       GATE                        SEDIMENT BUILD-UP
                                                     SEDIMENT BUILDUP
                     GATE VALVE

                                                                          INLET RAISER & SCREEN
               B. Pipe through base of dam wall

                                        Figure 6.4 Discharging of clarified water

6.8    Revegetation
       Successful revegetation is dependent upon adequate site preparation. This normally
requires several steps discussed earlier. Revegetation can itself be divided into several steps.

Assessment of Plant Nutrients and Conditions
       It is almost always more effective to select plant species which are suited to local
conditions than to undertake major amendment of the soil characteristics. This may not be
possible at difficult sites or tailings areas. An assessment of soil conditions will identify any
serious soil deficiencies and can indicate appropriate fertilizer applications. Then certain
measures of soil amendment must be carried out.
       Techniques that can enhance revegetation are discussed as follow.
(1) Soil Amendment without fertilizers
       Gypsum is used to condition heavy clay soils and reduce surface crusting on hard-
setting soils. The application of gypsum replaces sodium ions with calcium, which can
improve soil structure, increase water infiltration, aeration, reduce crusting and by leaching
reduce salinity. Application at the rate of 5 tonnes/hectare is normally sufficient to treat
surface crusting. Rates of 10 tonnes/hectare may be required for clayey subsoil. The
amendment effects of gypsum last several years during which re-vegetation will enable
increase in organic material to have a more sustained effect.
       Lime is used primarily to adjust pH but may also assist in improving soil structure.
Adjusting pH can significantly alter the availability of plant nutrients and toxins. Lime is
mostly applied as ground limestone and dolomite or agricultural lime. Hydrated (slaked)
lime is less common. Agricultural or coarsely crushed limestone and dolomite is slower
acting but is likely to have a more sustained effect on pH than rapidly acting slaked lime.
Subsequent application of lime may be required if sustained elevation of pH is required.
Hydrated lime will reduce the effect of nitrogenous fertilizers if applied at the same time. It
should be applied separately. The extent of pH adjustment rate of application will depend on
the extent of acidity, the soil type and the source of the limestone. As a guide, application of
agricultural lime at the rate of 2.5-3.5 tonnes/hectare will increase pH by approximately 0.5
units provided soil pH is not less than 5.0.
       Mulches are materials applied to the surface to enhance soil conditions for seed
germination and initial plant growth. Short-lived cover crops are also used as mulches. Use
of mulches can control erosion. Other benefits include retention of soil moisture, protection
of seedlings and modification of soil surface temperature. Application of most mulches is
limited to locations requiring rapid re-vegetation, special protection (embankment, etc.) or
where significant improvement of the soil or root medium is required (e.g. tailing dams
etc.). Hay and straw are commonly used for mulching broad scale areas. Application rates
vary between approximately 2.5 to 5.0 tonnes per hectare. A wide range of other organic

materials or agricultural wastes is suitable for mulching: their use is dependent on
availability and cost. Materials or wastes successfully used include stripped vegetation,
sawmill and wood processing wastes (chips and sawdust), sugar mill wastes (bagasse),
brushwood, and peanut and macadamia nut shells. Additional nitrogen may be required to
compensate for the nitrogen demand created when fresh mulch materials break down.
       Mechanical application of mulching materials can be by conventional agricultural
equipment (manure spreaders, etc.) or by specialized equipment. Specialized mulching
equipment applies mulching material (usually hay or straw) mixed with seed. The material
is broadcast either in dry from or as slurry with water or bitumen or specially formulated
soil-binding agents. Hydro mulching or similar techniques have the advantages of applying
mulches to relatively large but inaccessible areas, however the necessary equipment is
costly and is normally carried only by contractors. Cereal crops (cereal rye, oats, sorghum,
millet, etc.) can be grown to provide an in-situ mulch while permanent vegetation is being
established. In the mid-1970s, Quarry Industries Ltd. rehabilitated some of the more visible
overburden dumps of Stonefell Quarry by spraying with an emulsion containing wood fiber
and native plant seed. That method was known as hydro seeding. As a temporary measure in
1976, the most visible quarry faces were sprayed with bitumen to reduce visual impact. This
received a mixed reaction from residents and environmental groups.

(2) Using Fertilizers
       Requirements will vary widely according to the conditions and intended post-
mining land use. For agricultural uses fertilizer application will need to meet specific
requirements of proposed pasture or crops. Precise soil testing, multiple fertilizer
application and subsequent intensive maintenance may be required. Although native species
are adapted to low nutrient levels, improved growth can be achieved by modest application
of fertilizers. Species vary in their capacity to respond. Some are sensitive to increases in
levels of phosphorus and are likely to be adversely affected. Organic fertilizers (sewage
sludge, manure, blood and bone, etc.) are generally beneficial but often costly and difficult
to apply. Unlike most inorganic fertilizers they are beneficial both as fertilizers and as soil
       Application rates of inorganic fertilizers should be assessed according to soil
analysis. Where re-vegetation is with native species, relatively low application rates (250-
400 kg/hectare) of compound fertilizers have resulted in increased plant vigour.
Commercial inorganic fertilizers always contain one or more of the macronutrients (i.e.

nitrogen, phosphorus and potassium). They may also contain sulphur, calcium and
magnesium. The level of available trace elements (boron, copper, cobalt, iron manganese,
molybdenum, zinc, etc.) will vary with pH. There must be awareness of the potential for
excessive application of fertilizer to cause water pollution problems, particularly in areas
with sandy soils. Slow release fertilizers in the form of granules or tablets can be placed 10-
15 cm below or adjacent to individual tree seedlings at the time of planting. Direct contact
of fertilizers with seedling roots must be avoided.

6.9       Reclamation Opportunities
          It is a common perception that mining destroys land values. However, while mining
alters the shape of the land, it often provides a reconfigured piece of land that is of more
interest and value than it was in original state. Some of the land use possibilities are as
          1.    Crop-land
          2.    Pasture land or source of hay
          3.    Grazing land
          4.    Forestry use
          5.    Residential use
          6.    Industrial / commercial use
          7.    Recreation
          8.    Fish and wildlife habitats
          9.    Developed water resources, and
          10.   Undeveloped land or no current use.
          The planners should encompass the whole range of possibilities. Obviously, there
are several considerations. The overriding consideration is that the option be practical. The
second is that a demand should exist for the product of restoration. The third is that the
product should yield an economic or social return. Options for restoration of different types
of wasteland is shown in Table 6.2.

Table 6.2.Options for restoration of different types of wasteland
Old use            New use








Neglected          xx               x       x         x           x             x       ?               x                       x

Dry pit            x                x       x         x           x             x                       xx                      xx

Wet pit                                     x                     x                     xx              xx
Mine waste         x                x       ?         ?           x             x                       x                       x

Refuse tip         x                                              xx            xx                      x                       x

Housing                                     xx        xx          xx            x                                               x
Industry                                              xx          x             x                       x                       x

xx: likely new use; x : possible new use; ? : Possible new use depending on circumstance.

6.10      Species Selection
          Most mine site rehabilitation programs are directed toward the re-establishment of
native species. Selection of local species adapted to the prevailing climate and soil
conditions is recommended. Even with in a single species a wide range to tolerances has
been shown in response to site conditions. As a consequence, seeds of a species from one
location may not necessarily thrive when used elsewhere and collection of seed from areas
nearby the mine site is strongly recommended.
          The species selected will depend on the future land use of the area. There must be
awareness of the potential fire danger of re-vegetation areas. Where grazing, cropping, etc.
is the planned land use then the species planted will be those commonly used for pasture or
crops known to be successful on soil of similar texture, pH and fertility.
          Where the re-establishment of a diverse and permanent cover of local species is the
aim then the following methods of determining suitable species should be followed.
Observe plant species growing naturally on any old disturbed areas near the rehabilitation
site so that good colonizing and follow-up species can be identified.
          1. Observe the soil and drainage conditions to which the different local species are
             adapted and match them with the conditions on the mine site.

       2. Select plant species that produce sufficient viable seed, which can be harvested
       3. Consider habitat requirements where return of wildlife to the area is a significant
           element of post-mining land use.
       4. Consider planting legumes as they are often good colonizers and will improve
           soil fertility.
       Some species of grasses and legumes, and suggested crops for sand and slime
tailings in tropical climate are shown in Table 6.3 and Table 6.4.

Table 6.3. Species of grasses and legumes for tropical climate
                Common Name                        Botanical Name
Grasses         Sudan grass                        Sorghum sudanensis
                Reed Canary grass                  Phalos arundincea
                Signal grass                       Branchiaria brigantha or decumbens
                Para grass                         Branchairia mutica
                African star grass                 Cynodon plectostachyam
                Napier/elephant grass              Pennisetum purpureum
                Foxtail                            Setaria sphocelata or splenolide

Legumes         Common name varies                 Stylosanthes suyanensis
                  between countries
                Common name varies                 Centrosema pubescens
                  between countries
                Common name varies                 Crotolaria usaramulusio
                  between countries
                Common name varies                 Calopogonium desmoidum
                  between countries
                Common name varies                 Calopogonium caeruleum
                  between countries
                Crown vetch                        Coronilla viria
                Rice bean                          Phasedus calearatus
                Birds foot trefoil                 Lotus corniculatos
                Alfalfa/lucerne                    Medico sativa

          For the rehabilitation of mines in Myanmar, advice should be sought from ministries
for forest and agriculture that can assist in the procedures of selecting plant species for a
particular site.
          Table 6.4. Suggested crops for sand and slime tailings areas
                          Sand/silt areas       Slime areas

Grass/legume forage                             Grass/legume forage
Root crops: cassava, ground nuts, onions,       Paddy and upland rice
  yams, carrots                                 Maze
Melons                                          Soya beans
Leaf vegetables                                 Sorghum
Pepper                                          French beans
Cloves                                          Long beans
Tobacco- virginia type                          Tomatoes
Papaya                                          Chilly
Pomelo                                          Pineapple
Star fruit                                      Banana
Cashew                                          Tobacco - burley type
Citrus                                          Pomelo
Ipil-ipil (Leucaena leucocaphala)               Star fruit

          In Myanmar, there has been no record of mine rehabilitation; subsequently, no
specific indication of suitable species for mine rehabilitation is found. Trial of native
should be given priority. Depending on the diverse climatic, topographic and latitudinal
effects, Myanmar forest flora has been divided into eight major types. They are:
          1. Tidal forest
          2. Beach and dune forest
          3. Swamp forest
          4. Evergreen forest
          5. Mixed deciduous forest
          6. Dry forest

            7. Deciduous dipterocarp or Indaing forest, and
            8. Hill forest.
            Yet it is important to seek advice from respective departments.

6.11        Seed Collection and Extraction
            Before collecting seed it should be checked with relevant authorities to establish
what permits/licenses are required. It is unlikely the mine site will be uniform in terms of
aspect, soil types, etc. Seed will need to be collected from several areas so as to match site
conditions. Points to consider when collecting seed include:
   1. Locate collection areas and desired species before seed matures
   2. Avoid seed cases or fruit that show any signs of insect attack or fungal infestation.
   3. Collect seed only when it is mature. Differential ripening within one species or even
            a single plant may necessitate several visits.
   4. Where a plant has seed cases of varying ripeness, the oldest are those furthest from
            the growing tip. Check to ensure that cones or cases have not opened and seed
            already discharged.
   5. Avoid placing seed or seed cases in plastic bags. Use cloth or paper bags.
   6. Collection of seed can be speeded up by mechanical means, including modified
            vacuum cleaners and scoops fitted to the bull bars.
   7. Where air-drying is insufficient to open seed cases, heat extraction may be required.
            Oven temperatures of 60°C are suitable for Eucalypts, Casuarinas and Hakeas.
   8. If seed is purchased, ask for information as to the area from which the seed was
            collected and the date. Commercial seed should specify germination percentage and
            date tested.

6.12        Seed Storage
            Some considerations are:
       1.       Clean the seed to remove as much debris as possible before storage Failure to do
                so may result in fungal infection.
       2.       Label the seed clearly including species, date collected, location, etc.
       3.       Store the seed in dry; insect and vermin proof containers and dust with fungicide
                and insecticide powder.
       4.       Check longivity of seed before storing, most seed will keep for several years but
                some species cannot be stored beyond 1 or 2 months.

       5.      For most species storage at 1-4°C in airtight containers at less than 10 percent
               humidity will maintain viability of seed.
       6.      Tropical species may be killed at temperatures below 10°C.
       7.      Domancy in many plant species is affected by changes in light or temperature
               conditions and moisture content.

6.13        Seedbed Preparation
            Methods used for the preparation of the seedbed will depend on topography of the
site, the proposed land use, the extent of soil amelioration and fertilizer use, and the sowing
or planting technique proposed. Points to consider include:
   1. Prevention of compaction, crusting and subsequent erosion by avoiding disturbance
            to soils when wet and sticky or dry and powdery.
   2. Application of most fertilizer is best carried out during tillage of the seedbed.
            Nitrogenous fertilizers tend to dissipate and planting should follow immediately
            after application.
   3. Timing of seedbed preparation (and sowing) is often critical for successful
            establishment of vegetation. In most cases preparation and sowing should occur
            immediately prior to the onset of reliable rainfall.
   4. Ideally timing should enable seedbed preparation where soil moisture levels are
            adequate to avoid compaction or powdering but sufficient to induce germination and
            enable follow-up rains to sustain plant growth.
   5. In some areas soil temperatures also need to be considered. Local agricultural
            practices may provide a guide to the optimum period.
   6. To best use the often-limited period of suitable weather conditions, maximize use of
            mechanical equipment or when large areas are involved consider aerial applications.
   7. A variety of heavy-duty conventional agricultural equipment can be used for
            seedbed preparation. Disk harrows and chisel plows are both able to operate in stony
            soil conditions.
   8. Avoid over preparing the seedbed. A rough cloddy surface reduces runoff and
            provides better lodgment and protection for seeds and seedlings.
   9. When hand planting of seeds or seedling is proposed, site preparation may best be
            limited to deep ripping or minimal tillage.

6.14   Planting
       Several options are available where the re-establishment of native vegetation is
required. The planting method selected will depend on size and nature of the site and the
availability of plant material. Options and relevant considerations include:

(1) Direct Seeding
   1. Potentially a very economical method for re-vegetation but only where germination
       and seedling survival rates are sufficiently high.
   2. Advantages include low labor costs, random broadcast of seed and no check on
       growth rates through planting-out.
   3. Seed can be broadcast by hand, conventional or modified agricultural equipment, by
       air and hydro mulching, etc.
   4. Disadvantages include higher risk of failure through adverse climate conditions,
       competition from weeds, loss of seed by insect predation and low seed germination
       survival rates.
   5. Failure can result in loss of seed stocks and seasonal opportunity for field work.
       Before sowing, check to ensure pre-treatment of seed is not required to break
       dormancy by heating, soaking in hot water, acid or potassium nitrate.
   6. Seeding rates will depend on species, site conditions, and desired density of
       vegetation and viability of seed; test the seed to establish approximate viability
   7. Number of seeds per unit weight is highly variable.
   8. Ultimate survival rate of direct sowing is typically 1-5% for fine seed   species and
       5 -10% for coarse seed species. Assuming 75% seed viability rate, a rough guide to
       sowing rates is 100gms - 1.0 kg/hectare for fine seeded species, and 2.0 - 4.0
       kg/hectare for coarse seeded species. Specific site conditions may require much
       higher sowing rates.
   9. Bulking of fine seed with sand, sawdust, etc. will assist in even distribution.
       Pelleting larger native seed to assist in sowing and reduce wind blow has been
       successfully achieved.

(2) Planting Seedlings
       1. A reliable supplier of seedlings or the establishment of an on-site nursery is

      2. Advantages include efficient use of available seed, control over species mix and
          placement and no limitation on the species included in the re-vegetation
      3. Disadvantages include higher costs for planting and/or nursery operation or
          purchase of seedlings, check in growth rate at planting, need to pre-order or sow
          several months prior to anticipated use, longer planting time required and
          seedlings may deteriorate if planting is delayed.
      4. For sites requiring a large and/or sustained supply of seedlings, costs can be
          significantly reduced by establishing an on-site nursery. Specific skills or
          guidance are required to ensure appropriate nursery procedures that will provide
          a reliable seedling supply.
      5. Plant seedling along rip lines or on graded mounds if location is poorly drained.
      6. Unless irrigation can be provided, plant the seedling when soil is moist and prior
          to reliable rainfall, or in cold and wet areas, when water logging is least likely to
      7. Handle seedlings with care. Where seedlings are not in tubes or similar
          containers, ensure roots remain moist.
      8. Where seedlings are planted into an existing vegetation cover (cover crop or
          weed invasion etc.) clear the area adjacent to the seedling to reduce competition.
          This is particularly important if fertilizers are applied.
      9. Planting by hand need not be by shovel, mattock, etc. although this remains a
          widely recommended and reliable method.
      10. Devices to speed up planting include auger attachments for tractors, etc.,
          planting sticks suitable for sites with friable soils.

(3) Transplanting
   1. Transplanting of mature trees and shrubs is practicable for specific sites.
   2. Advantages include immediate effect and introduction of possible seed
      source.Disadvantage is high risk of costly failures.
   3. Where individual mature trees are required for the rehabilitation programme,
      transplanting should be completed while suitable earthmoving and lifting equipment
      is on site.

(4) Aftercare
   1. Success rate of all methods of planting will be reduced unless adequate aftercare is
   2. Staking and protection of individual trees is desirable but seldom practicable on a
       broad scale.
   3. Perimeter fencing will provide protection from browsing stock, vehicle and
       pedestrian trafficking. Temporary fences are unlikely to provide adequate protection
       for the required period.
   4. Perimeter fencing which incorporates wind-breaking materials will increase success
       of re-vegetation programmes in most circumstances.
   5. Avoid irrigating seeded areas unless irrigation can be maintained until reliable rains
       occur. Irrigation of seedlings should be progressively and slowly reduced to prevent
       over dependence of surface roots.
   6. Follow-up application of fertilizers, additional seedling or planting may be required.
   7. Damage and loss by insects and vermin are common, particularly where re-
       vegetation programmes provide herbage in otherwise sparse environments.
   Needles to say, technology improvement alone is not enough for operating an
environmentally sound mining project: government policy, regulations, and environmental
management systems (EMS) of organizations need to be properly linked.


                          MINE VENTILATION FOR B. TECH
                              CHAPTER 1 (GENERAL)


       Air is a mechanical mixture of a number of gases, each of which has different
physical and chemical properties. The main ingredients of pure air are oxygen and nitrogen,
but there are also present small percentages of rare gases such as argon, neon and helium
and a variable percentage of carbon dioxide. A good approximation of proportion is:

               Oxygen:        21percent by volume;
                              23 percent by mass.
               Nitrogen: etc, 79 percent by volume;
                              77 percent by mass.

       From the above it can be seen that oxygen is heavier than nitrogen – about 14%
heavier – but these gases nevertheless remain mixed in this proportion. If the heavier gases
were to settle out there would be a much higher concentration of oxygen and of carbon
dioxide, which is much heavier still, at the coast than at high laying areas such as the South
African gold fields, but this is not in the case.

        Air usually has a certain amount of water vapour associated with it. This water
vapour is conventionally spoken of as being part of the air, but actually it is not part of it
and rather constitutes an impurity similar to dust, odours and bacteria, which are also
normally present. Strictly speaking, it is wrong to say that air contains a certain amount of
water vapour, although this expression has been commonly accepted. It is more correct to
say that a given space contains a certain mass of air and a certain mass of water vapour. A
space of one cubic metre contains one cubic metre of air and one cubic metre of water
vapour. The water vapour can account for anything up to four percent of the mass of the

       The actual mass of a specific gas or vapour contained in a space of one cubic metre
is dependent on its temperature and on the pressure exerted on it. The pressure exerted on a
gas consists of atmospheric pressure and of additional pressures such as are caused by
fans and compressors.

       In scientific term pressure is a force exerted on an area. The unit of force, defined as
the product of mass and acceleration, is the Newton (N = kg.m/s2). The basic unit of
pressure in the international metric system – the Systeme International or SI – is thus the
Newton per square meter (N/m2). This unit has recently been named the Pascal (Pa).

        Atmospheric pressures and compressed air and water pressures and pressure
imparted to air by mine fans vary between some hundreds and many thousands of Pascals.
Because it is inconvenient to work with such large numbers, it is more common in mine
ventilation to express pressure in terms of kilopascals (kPa) where 1 kPa = 1000 Pa. A
pressure of one kilopascal is equivalent to a difference of just over 102 mm between the
water levels in a manometer.

         Another well-known unit of pressure, which is popular with meteorologists although
it is not part of the SI, is the bar. A bar (abbreviated from of the word barometer) is equal to

100,000 Newtons per square metre and is approximately equal to the mean barometric
pressure at sea level (101300 N/m2 = 101300 Pa = 101.3 kPa). No further mention will be
made of the bar as the South African mining industry has decided, since the publication of
the first edition of this book, to accept the Pascal and kilopascal as the units of pressure.

        Atmospheric pressure at any particular point is equivalent to the weight of the
atmosphere vertically above the point and it is consequently lower at high-lying places than
at low-lying places. For most practical purposes it is convenient to remember that the
atmospheric pressure at sea level is about 100 kilopascals and that it increases by 1
kilopascal for every 90 metres below sea level and decreases by       1 kilopascal for every
90 metres above sea level. Thus the pressure at 1700 metres above sea level (the shaft collar
elevation of the highest Witwatersrand mines) could be expected to be about 81 kPa while
the pressure at 1700 metres below sea level (the deepest point reached in a South African
gold mine at present) would be about 119 kPa.



        The temperature scale in the SI System is called CELSIUS scale (This is identical to
the old CENTIGRADE scale). It is based on the boiling and freezing points of pure water at
normal sea level pressure.

               Freezing point of pure water = 0 C
               Boiling point of pure water = 100 C

         About two hundred years ago the interesting discovery was made that: “ The same
rise of temperature produces in all gases the same increase in volume, provided the pressure
be kept constant”. This law is called Charles’ law.

        This statement is very nearly true for most of the common gases and for most
practical purposes it has been found that at constant pressure the volume of gas varies by 1 /
273 of its volume at freezing point (0 C) for every 1 C change in temperature.

        Thus 273 m3 of air for freezing point will change its volume, as follows (provided
that the pressure remains constant)

       At      1 C it will become 274 m3
               2 C it will become 275 m3
               10 C it will become 283 m3
               100 C it will become 373 m3
               1000 C it will become 1273 m3
               -10 C it will become 263 m3
               -100 C it will become 173 m3
               -273 C it will become 0 m3

         It is difficult to imagine that if a gas is made cold enough it will occupy no volume
at all. However, this anomaly can be explained by the assumption that it is the spaces
between the smallest particles of matter forming the gas that obey Charles’ law, and that at
– 273 C this matter in its most compact form will stay occupy a small volume.

        An interesting phenomenon that occurs at temperatures approaching – 273C is that
the electrical resistance of certain metals such as mercury, lead and tin drop to zero, there by
making them so-called superconductors. A current of electricity once started in a ring
made of a superconductor continues to flow indefinitely.

       The temperature –273 C is regarded as the lowest possible temperature that can be
obtained and it is therefore called the absolute zero temperature. The Celsius scale can
now be converted to an absolute scale, which is called the Kelvin scale.

         CELSIUS    KELVIN
Boiling water 100C    373K

 Melting ice       0C          273K

Absolute zero     -273C         0K

                    FIG. 1

       Absolute C = C + 273 = K

       The two scales are depicted in figure 1-1.

        Charles’ law can now be stated in a more convenient form, namely that the volume
of a gas at constant pressure is directly proportional to its absolute temperature, or

                       V1/V2 = T1/T2

       Where V represents volume and T represents absolute temperature.

Example 1: If 10,000 m3 of air at 20 C is heated to 27 C at constant pressure, what does
its volume become?

           10000/V2 = 273 + 20 / 273 + 27
                 V2 = 10000  300/293
                    = 10240 m3


       Boyle’s law states that the volume of a gas is inversely proportional to the absolute
pressure exerted on it, provided that the temperature remains constant.

           Thus PV = constant (where p represents pressure)
           or P1V1 = P2V2 at constant pressure

Example 2: A certain mass of air has a volume of 10000 m+ on surface where the
barometric pressure is 80 kPa. What would the volume of this same air be at the bottom of a
deep mine where the pressure is 120 kPa if the temperature stayed the same?

           80 100000 = 120  V2
                   V2 = 10000  80/120
                      = 6667 m3

Example 3: A kilogram of air on surface at the pressure of 82 kPa has a volume of 1.0 m 3.
What would the volume of a kilogram of compressed air at a gauge pressure of 560 kPa be
at the same temperature?

           The absolute pressure of the compressed air is 560 + 82 = 642 kPa
           82  1.0 = 642  V2
                V2 = 1.0  82 / 642
                    = 0.128 m3

      When both the temperature and the pressure of a gas vary, then both Charles’ law
and Boyle’s law must be used, as shown in the following example.

Example 4: 150 m3/s of air enters the downcast shaft of a mine at a pressure of 82 kPa and a temperature of 17 C. At the bottom of the
shaft the pressure is 115 kPa and the temperature 27 C. How much air will reach the bottom of the shaft if none of it is lost due to

           150 / (273+17) = V2 / (273 + 27)                        (Charles’ law)
                       V2 = 150 300 / 290
                           = 155 m3 /s
                   2  155 = 115  V2                              (Boyle’s law)
                       V2 = 82 155 / 115
                           = 111 m3/s

        Thus the volume is increased due to the increase of temperature, but reduced to a
greater extent due to the increase of pressure.

        Instead of doing calculation in two steps at shown above, it can be done in a single
step by means of a formula, which combines the two laws, viz.

                    P1 V1 / T1 = P2 V2 / T2
Hence            82  150 / 290 = 115  V2 / 300
and                         V2 = 82  150  300 / 115  290
                                = 111 m3 /s

        The value of PV/T is constant for a given mass of any particular gas, and when the
mass of the gas is one kilogram it can be said that PV/T = R or PV = RT, where R is the
figure which depends on a particular gas and is called the gas constant. The value of R for
dry air is 0.2871 if P is in kilopascals. T is in K and V, which is now the specific volume, is
in m3/kg.

        By means of this formula, it now becomes possible to calculate the density (mass
per cubic metre) or specific volume (volume of one kilogram) of dry air at any temperature
or pressure.

       Note that density = 1 / specific volume

Example 5 : What is the density of dry air at 90 kPa and 20 C?

       PV = RT
   90  V = 0.2871  293
         V = 0.2871  293 / 90
Density, w = 1/V = 90 / (293  0.2871)
          = 1.07 kg/m3



       The amount of water vapour that a space can obtain depends on the temperature in
that space. If the space contains as much water vapour as it can hold at the existing
temperature it is said to be saturated with water vapour at that temperature.

        If the small amount of water were introduced into a vacuum space, some of it would
evaporate and it would build up a certain pressure in the space, depending purely on the
ruling temperature. This pressure is called the vapour pressure. A few values are given in
Table 1, together with some other properties of water vapour.

     TABLE 1
                           Properties of Water Vapour

     Temperature        Vapour Pressure       Specific Volume        Latent Heat
         C                  kPa                                        kJ/kg
        -2                    0.55                 242                 2835
         0                   0.61                  206                 2835
       20                    2.34                    57.8              2454
      40                     7.38                   19.5               2406
      60                    19.9                      7.68             2358
      80                    47.3                      3.41             2308
     100                   101                        1.67             2257
     150                   476                        0.39             2114
     200                  1554                        0.13             1939
     250                  3970                        0.05             1714

        When there is sufficient water to allow the vapour pressure to build up to the full
amount applicable at the ruling temperature, the space is said to be saturated with water
vapour at the temperature. If insufficient water is present, the pressure does not build up to
this full pressure to the saturation vapour pressure at the same temperature is called the
relative humidity of the space.

        The ratio of the actual mass of water vapour to the mass of water vapour that could
be contained at saturation is called the percentage humidity. There is only a very small
difference in value between relative humidity and percentage humidity under any particular
set of conditions, but it is important that the definitions should be understood. The one is a
ratio of vapour pressures, while the other is a ratio of mass.

        In 1802, John Dalton showed that "the mass of vapour required to saturate a given
space at a given temperature, and consequently also the vapour pressure of a given liquid is
the same whether the vapour be left by itself or associated with other gases upon which it
has no chemical action” This is Dalton's Law which means, in other words, that the total
pressure of a mixture of gases and vapours is the sum of the partial pressures of the
constituents. Another way of expressing this is that each gas in a mixture behaves as if the
others were not present. It is as if each gas is unaware of the presence of the other gases.
        The boiling point of a liquid is the temperature at which the vapour pressure of the
liquid is equal to the external pressure on the liquid surface. Thus water at a pressure of 101
kPa (sea level atmospheric pressure) will boil at 100C, while at a pressure of 3970 kPa
(inside a boiler) it will boil at 250C and at 0.55 kPa it will boil at -2C. This is below the
normal freezing point of water, and under these conditions the ice sublimes into vapour, i.e.
it changes directly from the solid to gaseous state without passing through the liquid state.

        If air containing some water vapour is cooled, a temperature will be reached at
which it will no longer be able to hold all the vapour and the vapour will start to condense
just as dew condenses from the atmosphere on a cool night. This temperature is called the
dew - point temperature. It can be obtained from psychrometric charts if the wet and dry
bulb temperatures and either the relative humidity or the barometric pressure are known.
Alternatively it can be measured directly by cooling an air-vapour mixture and measuring
its temperature at the moment when dew starts forming on a polished metal surface.


        It has previously been explained how the density of dry air can be calculated.
Calculating the density of humid air is rather more complicated and is normally avoided by
looking up the answer in density tables or charts. However, it is useful to know the
principles on which the calculation is based.

         In Example 5 it was shown that the density of dry air at 90 kPa and 20C was 1.07

Example 6: What is the density of humid air at 90 kPa and 20C?

       It is impossible to give an answer to this question unless more information is
supplied to indicate how humid the air is. Either the wet bulb temperature, the dew point
temperature or the relative humidity must be stated.

(a) Let us assume that the air is saturated with water vapour. From a complete set of steam
tables (of which Table 1, Note 4, is a very abbreviated example) it can be determined that
the vapour pressure of water at 20C is 2.34 kPa, and that a cubic metre of saturated vapour
at 20C has a mass of 0.017 kg (i..e.1/57.8)

         The partial pressure of the air is thus 90-2.34 = 87.66 kPa (See Note 4).
         Now using this pressure:

                PV = RT
          87.66  V = 0.2871  293
                 w  = 1/V = 87.66 / (0.2871  293) = 1.042 kg/m3

         This is the mass of air in one cubic metre of the mixture.

         The total mass of air plus water vapour in a cubic metre of the mixture is thus

                1.042 + 0.017 =1.059 kg

       One cubic metre of dry air and saturated water vapour thus has less mass than one
cubic metre of dry air at the same temperature and pressure.

(b) Let us assume that the relative humidity of the air is 50 per cent.
The partial pressure of the water vapour is now 50 per cent of 2.34 kPa = 1.17 kPa
The partial pressure of the air is thus 90 –1.17 = 88.83 kPa

         Calculating by the same method as before, we now find that the mass of air in a
cubic metre is 1.056 kg and the mass of water vapour approximately 50 per cent of 0.017
i.e. 0.009 kg.

        The total mass in a cubic metre of mixture with 50 percent of relative humidity is
thus 1.056 + 0.009 =1.065 kg.

       This is seen to be greater than in the case of a saturated mixture.

        (c) If the dew point temperature is given, the saturated vapour pressure at the dew
point is looked up in the tables and the calculation is then continued as in (a) above.

        In various tables and charts the humidity of air will be found in different units. In
mine ventilation work it is most convenient to work in units of grams of vapour per
kilogram of dry air because as the air travels through the mine its temperature and pressure
continually change and water vapour is added to or removed from it. Both the volume of the
air and the mass of the mixture are therefore continually changing, but the mass of dry air
stays constant.



        Having discussed some of the basic physical properties of air, we are now in
position to start talking about how this air is used for ventilating a mine.

      Air is required in amine for four main reasons - to supply oxygen for breathing to
remove heat and to dilute and remove dust and gases.

Oxygen for Breathing

         When a man breathes, a certain amount of oxygen is retained in his lungs and most
of it is replaced by carbon dioxide. The amount of air breathed by a man and the amount of
oxygen retained depend on his state of muscular activity. A man asleep or at rest does not
breathe as rapidly or as deeply as a man performing hard manual labour.

         A man doing hard work absorbs about 3 litres of oxygen per minute and a sleeping
man only about one-tenth as much. One litre of fresh air contains about 0.2 litres of oxygen,
and the human lungs are not able to extract all this oxygen. If the oxygen content of the air
is reduced from 21 per cent to 17 per cent a man can still work comfortably but will breathe
a little faster and deeper. Symptoms like dizziness; buzzing in the ears, rapid heart action
and headaches start appearing at about 14 per cent and death occurs at about 6 per cent.

        The percentages given above apply to the case where the oxygen concentration is
reduced without the carbon dioxide concentration being increased- for example when the air
is diluted by methane gas or when oxygen is absorbed due to the oxidation of metals and

minerals. However, when oxygen is depleted purely by breathing, it is replaced by carbon
dioxide and this also has an affect on humans.

       Carbon dioxide acts as a stimulus, which regulates the breathing of a person. When a
man increases his rate of work the amount of carbon dioxide released in the alveoli of his
lungs immediately increases. The concentration increases and this stimulates faster and
deeper breathing. This principle is applied in resuscitating persons who have been
suffocated or drowned, by supplying them with a mixture of oxygen and carbon dioxide.

        When the percentage of carbon dioxide in atmospheric air, which is normally only
about 0.03 per cent, is increased to 3 per cent, the rate of breathing is doubled. When it is
increased to 5 per cent, the rate of breathing is quadrupled and becomes laboured. Higher
percentages cause violent panting and severe throbbing in the head, but the concentration
has to be very high, if sufficient oxygen is still present in the air, before death will occur.

       When people breathe the air in a confined space, the oxygen concentration is
decreased while the carbon dioxide concentration is increased at the same time. In South
African mining regulations it was once laid down that about 0.015 m3/s of fresh air had to
be supplied for each person working underground at any one time, that the concentration of
carbon dioxide could not be allowed to exceed 0.2 per cent. A person doing hard muscular
work exhales at a rate of more than 0.05 litres of carbon dioxide per second, and this will
cause a concentration of 0.3 per cent in 0.015 m3/s of air. Apparently it was assumed that
the average underground employee did not over-indulge in hard muscular exertion!!
However, even if they did, 0.015 m3/s of fresh air would be ample to provide excellent
conditions. With 0.005 m3/s per person the oxygen concentration would only be reduced by
1 per cent, the carbon dioxide would become 1 per cent, and these are both still very safe

        The average volume of fresh air actually supplied per person in South American
gold mines is approximately 0.11 m3/s - about 0.08 m3/s in the shallower mines and about
0.14 m3/s in the deeper mines. Thus there should not be much danger of an oxygen
deficiency in general mine air, but unventilated dead ends should always be regarded as
dangerous and should never be entered without a flame safety lamp. Apart from indicating
the presence of explosive gases, such a lamp also indicates oxygen deficiency because the
flame is extinguished if there is less than about 17 per cent oxygen in the air.


                           CHAPTER 2 (HEAT AND HUMIDITY)


        The ventilating air that circulates through a mine is heated by various agencies with
the result that working places in deep mines are usually hot. One of the main purposes of
mine ventilation is to keep the working places as cool as possible.

         The principal sources of heat affecting the ventilating air are auto-compression and
heat from rock, water, machinery and lighting, men, explosives, compressed air mains and
electric cables, oxidation of minerals and timber, rock movements and mine fires.
         Before discussing theses various sources of heat, it is necessary that we should know
a little more about heat itself.

       Heat is a form of energy. The Law of Conservation of Energy states that energy
can never be destroyed. It can only be changed from one form to another. Some common
forms of energy are potential (due to position, e.g. a book on your desk which contains the
energy to fall to the ground if allowed to), kinetic (due to motion), electrical,
electromagnetic, sound, light, chemical and heat.

       Heat is one of the commonest forms of energy. It is produced by chemical action
such as the oxidation of metals (slow) or the burning of a match (fast), and also by friction
and impact, and in many other ways.

         Heat energy can be converted to mechanical energy by producing steam, which
drives an engine and this engine can drive a generator, which produces electricity, or
alternatively, it can hoist rock out of a mine, thereby being converted into potential energy
in the rock. Heat can also be converted directly to electricity by means of a thermocouple,
or to light by making a piece of metal red hot or while hot.

        Conversely electricity can be converted directly to heat in a heater, and mechanical
energy is converted to heat by friction. We warm our hands by rubbing them together, and
when the breaks of a fast-moving car are applied suddenly the tyres get so hot that they start
smoking. Even the best roller bearings are not quite frictionless and consequently become
hot if not cooled by means of air or oil. Impact is the worst kind of friction and normally all
the mechanical energy involved is converted into heat. When an anvil is continually struck
by means of a hammer, both the anvil and hammer become hot. Even the small amount of
energy, which goes into producing noise, is eventually converted into heat.

        Heat must not be confused with temperature. Heat is energy while temperature is
purely a state. One object can be at a much higher temperature than another and yet contain
much less heat. For example, a bath full of water at a temperature of 30C contains much
more heat than a cupful of hot coffee at a temperature of 70C. Temperature is measured by
means of a thermometer graduated according to one of the scales described in Note 2.
Temperature can also be felt by means of the hands or any other part of the body. An object
at a higher temperature feels hotter than one at a lower temperature and heat always flows
from a higher to a lower temperature never the other way around.
        Heat, however, is not easy to measure and in practice it is usually calculated from
other measurements, which include temperature. In order to specify an amount of heat we

need to have a unit just as we have units of length (millimetres, meters, kilometres), units of
mass (grams, kilograms) and also units of time, electricity, compressed air, etc.

        The definitions of units are always made as simple as possible so that scientists in
different parts of the world can easily check on the reliability of their meters and other

       The unit of energy and therefore heat is the Joule, which is defined as being one
Newton meter. A watt is equal to a joule per second and thus, for example, a 1 kW electric
heater will produce 1 kilo joule of heat every second. Prior to the adoption of the SI system
of units, the unit of heat was defined in terms of the energy needed to increase the
temperature of a unit mass of water by one degree. Thus

        (a) In the British of F.P.S (Foot, Pound, Second) system, a British Thermal Unit
(Btu) is the amount of heat required to raise the temperature of one pound of water by one
degree Fahrenheit.

       (b) In the Metric or C.G.S (Centimetre, Gram, Second) system, a Calorie is the
amount of heat required to raise the temperature of one gram of water by one degree
       (The calories, which are quoted in dietetics, are really kilocalories.)

       The following are some other useful basic definitions:

        Force is that which changes or tends to change the state of rest or of uniform motion
of a body.
        The Mass of a body is the quantity of matter in the body and it is always constant.
        The Weight of a body is that force with which the earth attracts the body. It varies
inversely as the square of the distance between the body and the centre of the earth.
        Work is performed when a force acts through a distance. When a force P acts
through a distance D the work done is equal to PD.
        The unit of work is also the joule and it represents the amount of work done when a
force of one newton moves its point of application by one metre in the direction of the
        Power is the rate of doing work. The unit of power is the watt and it is defined as a
rate of 1 Newton metre per second or 1 joule per second.
        Energy is the capacity of doing work and it is measured in the same units as a work,
namely joules.
        It was Dr. Joule who first proved that work is converted to heat; the expenditure of a
certain amount of work always results in the production of the same amount of heat. This
relationship is such that 4.187 kJ raises the temperature of 1 kg of water by 1ºC, or 4.187 J
raises the temperature of 1g of water by 1ºC; thus
        1 calorie = 4.187 J



       Equal masses of different substances require different amounts of heat to raise their
temperatures by the same amount. We have seen that 4.187 kJ is required to raise the

temperature of 1 kg of water by 1ºC. It has also been found by experiment that only 0.126
kJ is required to heat 1 kg of gold through 1ºC, while 1.76 kJ is required to heat 1kg of
wood through 1ºC, etc.

        These are rather long and clumsy statements, and it has been found convenient to
use the concept of thermal capacity, which is defined as follows: -

       The thermal capacity of a substance is the amount of heat required to raise the
temperature of a mass of 1 kg of the substance by 1ºC.

       By this definition the thermal capacity of gold is 0.126 kJ/kgºC and that of wood is
1.76 kJ/kgºC.

       Table 2 is a list of the thermal capacities of various common substances.

        It will be noticed that two different values have been given foe each of the gases.
The reason for this is that the temperature of a gas is affected by the pressure applied to it.
When a gas is compressed it becomes hot. This is the reason why a bicycle pump becomes
hot and also why the air on the delivery side of a compressor is hot. Conversely, a gas cools
when its pressure is reduced while it is doing work. This is the reason why the exhaust air of
a rock drill is cold.

         Now, when 0.712 kJ of heat is added to 1 kg of air confined in a cylinder, the
temperature of the air will increase by 1ºC but the same time the air will want to expand and
because it is not allowed to do so the pressure inside the cylinder will increase. If the
pressure is allowed to return to normal by allowing the gas to expand against a piston so that
it does work, the temperature will drop slightly and another 0.293 kJ has to be added to the
air to get it back to its previous temperature. For this reason we say that the thermal capacity
of air is 0.172 kJ/kgºC at constant volume and 1.005 kJ/kgºC at constant pressure. The ratio
of these two figures (which for air is equal to 1.4) is also a very important number in the
study of thermodynamics, but consideration of it is beyond the scope of these notes.

                                            TABLE 2

                                              Thermal capacity kJ/kg ºC
              Substance            At Constant Pressure    At Constant Volume
                 Air                        1.005                    0.712
            Carbon Dioxide                  0.837                    0.628
             Water Vapour                   1.884                    1.382
              Hydrogen                     14.236                   10.045

               Mercury                                   0.126
                Petrol                                   2.094
                Water                                    4.187
                 Gold                                    0.126
                 Lead                                    0.126
                Copper                                   0.377
                 Iron                                    0.461
                 Glass                                   0.753
               Quartzite                                 0.837
                 Coal                                    1.005
                Wood                                     1.758
                  Ice                                    2.094

        From the definition of the term thermal capacity it appears to be a simple matter to
calculate the heat content of any substance if its thermal capacity, its mass and its
temperature are known. Actually, this is not the case because to find the total heat content
one would have to calculate from the basis of the absolute zero temperature (-273ºC) and
unfortunately the thermal capacity of a substance is not constant at all temperatures. Also,
the substance would change its state at different temperatures, such as water vapour
becoming first water and then ice at lower temperatures. However, it is nevertheless easy to
calculate the additional heat required to change the temperature of a substance, as shown in
the following examples.

Example 7: How much heat is required to raise the temperature of 0.1 m3 of bath water
from 15ºC to 45ºC? (It should be noted that the density of water is 1000 kg/m3.)

          Heat required      = mass  temperature rise  thermal capacity
                             = (0.1 1000)  (45-15)  4.187
                             = 100  30  4.187
                             = 12561 kJ.

Example 8: At what rate is heat removed from 3 m3/s of dry air at sea level density (1.2
kg/m3), which enters a cooler at 35ºC and leaves it at 20ºC?

          Heat removed       = (3  1.2)  (35-20) 1.005
                             = 3.6  15  1.005
                             = 54.3 kilo joules per second (kJ/s)
                             = 54.3 kW

Example 9: If 300 000 tons of rock are broken per 30-day month in a mine at a depth where
the average virgin rock temperature is 40ºC and by the time this rock reaches surface in a
skip it has been cooled to 20ºC, what is the average rate at which heat is released by the
broken rock?

     If this mine is ventilated by means of 700 m3/s of air at a density of 1.2 kg/m3, by
how much could this heat increase the temperature of the air?

         Heat removed by rock
         = 300 000  1 000  (40-20)  0.837 kJ/month
         = (300 000  1 000  20  0.837) / (30  24  60 60) kJ/s
         = 1938 kJ/s
         = 1938 kW

           Heating 700 m3/s of dry air by 1.C requires
           700  1.2  1 1.005
           = 844 kJ/s
           = 844 kW

       The air temperature would thus increase by (1938/844)  1 = 2.3ºC.

       The last two examples have been calculated on the assumption that the air contains
no water vapour. In practice this is not the case, and it will be shown later that evaporation
and the presence of water vapour have a considerable effect on the answers.



       The solid crust of the earth is on average only 50 kilometres thick. It is called the
Lithosphere and has an average density about 2½ times that of water. Below this is the
Barysphere, which is about twice as dense and which is in a very hot state estimated at
about 55000 ºC at the centre. There is some doubt as to whether it is kept from melting by
reason of the great pressure of the overlying rocks.

        The temperature of the earth, surface depends mainly on atmospheric conditions. In
Polar Regions it is below freezing point, whereas in equatorial desert regions it can be very
hot during the day and quite cool at night. However, if a thermometer is buried a few meters
below the surface it will be found that there is hardly any change between day and night
temperatures and only a relatively small change from month to month. The following table
shows average temperatures taken at the Potchefstroom Agricultural College over a number
of years.

                                           TABLE 3

                                                  Average Temperature ºC
        Depth below Surface
                                            summer                      winter
                 1                            23.5                       13.4
                 2                            21.9                       15.4
                 3                            20.9                       16.3

        When these figures are graphed and the graphs extended, the conclusion is reached
that at a depth of about 6 metres the temperatures of the soil must be about 19ºC both in
summer and in winter.

       In mining practice, when we walk about the surface rock temperature we mean the
temperature at a depth of about 15 metres. On the Witwatersrand this is accepted to be
18.3ºC and in the Orange Free State 19.4ºC.

        With such low temperatures existing on surface and a temperature of 55000ºC at the
centre of the earth - 6400 kilometres down - the rock temperature would increase by about
1ºC with every 120 metre increase in depth, if the increase were to be uniform. In practice,
however, the increase is not uniform. It depends on the heat conductivity of the rock and
consequently varies from about 1ºC per 40m in Karroo shales and even less than 40m in
some other rocks, to 1ºC per 110m in the quartzites of the Witwatersrand system.

        The variation in the Geothermal Gradient can be explained most simply by means
of the following example. If one placed a 20 cm thick block of wood and a 20 cm thick slab
of copper on a stove which had a temperature of say 250ºC, then after a time one would still
be able to touch the top of the block of wood because, being such a bad conductor of heat,
its temperature would have increased to only say 50ºC, while the copper, being a very good
conductor of heat, would be nearly as hot as the stove - say 230ºC. Thus, the temperature
gradient across the 20 cm of wood would be          250 - 50 = 200ºC or 10ºC per cm, while
the temperature gradient in the copper would be 250 - 230 = 20ºC or only 1ºC per cm, i.e.
the temperature gradient across a good constructor is lower than across a bad conductor.

        Fortunately the quartzite of the Witwatersrand system are good conductors and
therefore the rock temperatures does not increase rapidly with depth in the Central Rand
mines where this is the only type of the rock present. In the Free State and some other areas,
however, the Witwatersrand System is overlain by varying thickness of Karroo shales and
Ventersdorp lavas which are not such good conductors, but even here we are quite fortunate
compared to some foreign mining areas, as shown by the following examples:

                                           TABLE 4

                                          Rock Temperatures .C at Various
                                               Depths Below Surface
            Mining Area
                                      1000 m            1500 m            2500 m
            Central Rand                 25                30                40
          Klerksdrop Area                32                37                47
          Orange Free State              35                41                54
             Kolar-India                 40                46                59
          Bralorne-Canada                35                50                 -

        These temperatures must not be regarded as anything but approximate averages
because temperatures vary from mine to mine in the same field, and even between different
positions at the same depth in one and the same mine.

        Now this high temperature of the rock is one of the main sources of heat in deep
mines. Not only does the rock which is broken every day become cooled by the ventilating
air as indicated in Example 9 (Note 8), but heat also flows into the air from the unlimited
masses of rock surrounding the underground airways and working places, and this is where
our good fortune turns into misfortune. The same property of high conductivity of the
Witwatersrand Quartzites which results in relatively low rock temperatures at depth, also
results in a very rapid inflow of heat when cool air is passed through airways. It is before
necessary to maintain a large volume flow of air through an airway in order to keep the
increase in air temperatures within reasonable limits.

        Another effect of high rock temperatures is that any fissure water issuing from the
rock is also at a high temperature and tends to heat the ventilating air. Actually an easy way
of determining the temperature of the virgin rock is to measure the temperature of any issue
of fissure water because, except in rare circumstances, these two temperatures are identical.

       These are various ways of reducing the flow of heat from the rock into the air, but
these will be discussed at a later stage.



        In Note 8 certain calculations were made to show the effect of heat on air, but for the
sake of simplicity it was assumed that the air contained no water vapour. Unfortunately
natural air always does contain water vapour and in mine not only the water vapour content
but also the heat content and the temperature and pressure of the air are continually
changing. The study of these changes in the properties of the air is called Psychrometry and
it involves very complicated calculations. However, these calculations have been greatly
simplified by the production of various set of tables and charts such as the publication
"Psychrometry and Psychrometric Charts" complied by A.W.T Barenbrug and published by
the Chamber of Mine of South Africa.

          It is necessary to know three different properties of an air-vapour mixture before its
psychrometric state can be fully determined - e.g. its pressure and temperature and one other
property, which indicates either how much water vapour or how much heat it contains. For
this last purpose it is generally most convenient to measure the wet bulb temperature.

        The true wet bulb temperature is the temperature indicated by a thermometer, the
bulb of which is covered with a thin layer of water (usually contained in a muslin wick) and
which has an air velocity at least 3 metres per second passing over it. Radiation shielding is
also required when in sunshine or near surfaces, but not when the air and the surroundings
are at similar temperatures. It must be clearly understood that a temperature measured in
this way is not the temperature of the air and cannot even be correctly called the wet bulb
temperature of the air - it is purely the temperature of the wet thermometer bulb when
exposed to air in certain state. In order to understand this statement, it is necessary first to
understand the concept of latent heat.

       Ice, water and steam (or water vapour) are different forms or states of one and the
same substance, but they are vastly different in appearance and in physical properties. These

differences are due to the fact that the molecules of H2O in the different states contain
different amounts of energy.

        If heat is added to a mass of 1 kg of ice, which has been cooled to below freezing
point, the temperature of the ice will increase by 1ºC for every 2.094 kJ of heat, which is
added (see Note 8), until it reaches the temperature of 0ºC. When more heat is added some
of the ice melts into water but the temperature of both the water and the ice remains at 0ºC
until 335 kJ has been added, at which stage all the ice will have been turned into water at
0ºC. With further heating, the temperature of the water will increase by 1ºC for every 4.187
kJ of heat which is added. When 418.7 kJ have been added the water will be at 100ºC (at
sea level). When more heat is added the temperature does not increase but some of the water
evaporates to become steam. When about 2260 kJ have been added, all the water will have
been turned into steam at 100ºC. After this the temperature of the steam is increased by 1ºC
for every 1.88 kJ that is added.

        Now if this steam is cooled, exactly the reverse will occur. About 2260 kJ have to
remove to turn 1 kg of steam at 100ºC into water at the same temperature, and 335 kJ have
to be removed from 1kg of water at 0ºC to turn it into ice at 0ºC. This heat which has to be
added or removed in order to change the state of a substance without affecting its
temperature, it is called latent heat, as opposed to sensible heat which changes the
temperature of the substance. The world latent means concealed or dormant while the world
sensible means perceptible to the senses or that, which can be felt. There are not really two
kinds of heat, but the heat is used in two different ways - either to raise the temperature or to
change the state of a substance and this reason it is given two different names.

       Thus we can now say that:
       The thermal capacity of water vapour is 1.884 kJ/kgºC.
       The latent heat of condensation of water vapour is 2260 kJ/kg at 100ºC. At normal
air temperatures this value is nearer to 2460 kJ/kg.
       The thermal capacity of water is 4.187 kJ/kgºC.
       The latent heat of fusion of ice is 335 kJ/kg.
       The thermal capacity of ice is 2.094 kJ/kgºC.
       The thermal capacity of dry air at constant pressure is 1.005 kJ/kgºC.

        All the above figures are affected to some extent by temperature or pressure and
they are therefore not really constants. However, every small variation is taken into account
when doing very accurate calculations and this has also been done by the compilers of the
various psychrometric charts and tables.

       To come back to the wet bulb thermometer, what happens to it is the following: -

         Firstly, if the air is saturated with water vapour, the wick on the wet bulb and the
water in it will assume the temperature of the air and the thermometer will indicate exactly
the same temperature as the dry bulb thermometer -- say 25ºC. The only difference will be
that it will take a little longer than dry bulb thermometer to assume this temperature.

        Secondly, if the air is not saturated but has a relative humidity of say only 50 per
cent, a much more complicated process takes place. Water from the wet bulb immediately
starts evaporating into the air. This evaporating water requires considerable amounts of heat
to turn it from water into water vapour and the only substances it can draw this heat from

are the remaining water, the wick, the thermometer bulb and the surrounding air. In this
process the temperature of the remaining water is thus lowered, but it is not lowered
indefinitely because, as soon as it becomes cooler than the air, heat also starts flowing from
the air to the water. Thus there are two opposing actions taking place at the same time --
heat leaves the water due to evaporation and enters it from the air due to conduction. After a
time these two actions must find their point of balance at which as much heat enters the wet
bulb as is leaving it. The temperature of the wet bulb now becomes steady and can be read
off. For example, assuming a barometric pressure of 85 kPa, a dry bulb temperature of 25ºC
and relative humidity of 50 percent, the wet bulb temperature will be 17.6ºC. If the relative
humidity had been 80 per cent, the wet bulb temperature would have been 22.3ºC while at
20 per cent relative humidity it would have been as low as 11.7ºC.

        This effect of unsaturated air on the wet bulb thermometer is exactly the same effect
we obtain from a canvas water bag, which supplies cool drinking water on a hot day if we
hang it in the shade with a breeze blowing over it. As a rule such a bag would cool water to
a much greater extent in Klersdrop than in Durban because the relative humidity at the
former is usually much lower than at the latter. On the other hand an impervious plastic
water bag would not cool the water at all because no water would leak through it to
evaporate from the outside. Thus after sufficient exposure the water in a plastic water bag
will assume the (dry bulb) temperature of the air while the water in a canvas bag would
assume the temperature of the wet bulb thermometer.



        The air temperature and the temperature of the wet bulb thermometer -- commonly
called the dry and the wet bulb temperatures of the air -- are measured by means of a
whirling hygrometer which consists of two thermometers installed in a frame with a
handle with a handle by which it can be whirled rapidly. The bulb of the one thermometer is
covered with a wet cotton wick, one end of which is immersed in a water reservoir.

       Barometric pressures are measured by means of an aneroid or a mercury barometer.
Once the wet and dry bulb temperatures and barometric pressure at any place have been
determined it is possible to calculate or determine from charts all the psychrometric data
concerning the air in that place. The calculation of the density of the air has already been
explained in Note 5. Taking the same figures as were used in Example 6 (b) in that Note, it
will now be shown how the heat content of the air is calculated.

       In this case, we were given a barometric pressure of 90 kPa, a dry bulb temperature
of 20ºC and a relative humidity of 50 per cent, which corresponds to a wet bulb temperature
of 13.5ºC. It was then shown that a cubic metre of this air contained 1.056 kg of air and
0.009 kg of water vapour, giving a density of 1.065 kg/m3.

        The specific volume of this air is 1/1.065 = 0.939 m3/kg which means that 0.939 m3
of the air vapour mixture has a mass of one kilogram.

       The specific volume of the dry air is 1/1.056 = 0.948 m3/kg which means that 0.948
m of the mixture contains one kilogram of air plus some water vapour - in this case 0.948 ×
0.009 = 0.00853 kg of water vapour or 8.53 grams of water vapour.

        The total heat content of this 0.948 m3 of mixture is the total amount of heat required
to heat 1 kg of air and 0.00853 kg of water vapour from absolute zero                 (-273ºC)
through all its phases to the existing temperatures. However, it is inconvenient to calculate
this value and because in practical mine ventilation we never work with temperatures below
the Celsius zero, we usually find it more convenient to work with the heat required to raise
the temperature from 0ºC to the existing temperature.
        Now assuming that we have 1 kg of air plus 0.0085 kg of water both at 0ºC and that
these have be heated up to 20ºC dry bulb and 13.5ºC wet bulb, the following amounts of
heat are required if the specific heat and latent heat constants given in Note 10 are used:

(a) To heat the air from 0ºC to 20ºC requires 1 × 1.005 × 20 - -- -- --              20.10
(b) To heat the water from 0ºC to 13.5ºC requires 0.00853 × 13.5 × 4.187               0.48
(c) To evaporate the water at 13.5ºC requires 0.00853 × 2460    -- -- --              20.98
(d) To superheat the water vapour from 13.5 to 20ºC requires
     0.00853 ×1.884 × 6.5 -- -- -- - - -- -- -- -- -- -- -- -- --                     0.10

        Total Heat (or Enthalpy) above 0ºC - -- -- -- -- -- -- -- -- --              41.66

         Item (a) is called the sensible heat of the air, (b) is the sensible heat of the liquid, (c)
is the latent heat of the vapour and (d) is the sensible heat of the vapour.

        Instead of doing this long calculation, one can look up the answer directly from the
chart. The answer will probably be slightly different because of two reasons. On the other
hand the charts should be more correct because they are based on calculations using more
exact values for thermal capacity and latent heat. On the other hand they can be less
accurate because blocks for different colours may not have been exactly superimposed
during the printing of the charts.

        Instead of Total Heat (or Enthalpy) some charts give Sigma heat. Sigma                  ( ∑,
the Greek letter S ) heat is a term used for the total heat of the air less heat of the liquid. The
calculation is done in exactly in the same way as above, except that item (b), which is the
heat of the liquid water, is omitted. Thus the Sigma heat of the air at 13.5 /20ºC and 90 kPa
is 20.10 + 20.98 + 0.10 = 41.18 kJ/kg. This value is more conveniently shown in charts and
tables because it varies only with the wet bulb temperature and the barometric pressure and
is not affected by the dry bulb temperature. Total heat (enthalpy) on the other hand does
vary slightly with the dry bulb temperature.

        Whether the total heat or sigma heat values should be used in calculations which are
carried out in order to determine how much heat has been added to or removed from air
during heating or cooling processes, depends on the original temperature of the water which
is evaporated or on the final temperature of the water which is condensed in then process.
The true answer usually lies somewhere between the two but in most practical applications
sigma heat values, apart from being easier to apply, also give the more correct result. It is
only in calculations involving no change in moisture content of the air that the use of
enthalpy gives more accurate answers than the use of sigma heat.



                                            TABLE 5

                           Temperature ºC
   Line       Barometric                       Sigma     Moisture    Relative     Specific
  Number       Pressure                         Heat     Content     Humidity     Volume
                 kPa         Wet     Dry       kJ/kg      g/kg         %           m3/kg

      1            85         15      15        46.8       12.8         100          0.995
      2            85         16      16        49.9       13.6         100          0.995
      3            85         25      25        84.1       24.2         100          1.047
      4            85         26      26        89.0       25.7         100          1.053
      5            85         25      35        84.1       20.0          46          1.075
      6            85         25      37        84.1       19.1          39          1.080
      7           100         15      15        41.6       10.8         100          0.842
      8           100         16      16        44.1       11.5         100          0.846
      9           100         25      25        74.9       20.3         100          0.885
     10           100         26      26        78.8       21.5         100          0.889
     11           100         37      37       137.6       41.6         100          0.950
     12           100         38      38       144.6       44.2         100          0.957
     13           100         25      30        74.9       18.3          66          0.896
     14           100         25      37        74.9       15.4          37          0.912
     15          102.5       16.1    17.1       43.8       10.8          88          0.829
     16          102.5       25.9    27.1       77.1       20.3          90          0.870
     17          102.5       37.7    39.1      139.8       41.6          90          0.934
     18           115         15      15        38.9        9.5         100          0.731
     19           115         16      16        41.3       10.2         100          0.734
     20           115         16     17.6       41.3        9.5          84          0.737
     21           115         25      25        69.3       18.0         100          0.767
     22           115         26      26        72.9       19.1         100          0.771
     23           115         26     28.7       72.9       18.0          80          0.776
     24           115         37      37       126.0       36.8         100          0.821
     25           115         38      38       132.0       38.9         100          0.827
     26           115         38     42.9      132.0       36.8          72          0.836
     27           115         25      35        69.3       13.9          43          0.787
     28           115         25      37        69.3       13.1          37          0.791

       Table 5 gives a number of values, which were looked up in Barenbrug's
psychrometric charts.

       Some every interesting facts emerge from a detailed study of these figures.

        (a) When lines 1, 2, 3 and 4 are compared it is seen that they all represent saturated
air at the same pressure of 85 kPa. As the temperature increases the heat content and the
moisture content increase and the specific volume also increases - i.e. the density decreases.

        (b) When lines 3, 5 and 6 are compared, it is seen that the barometric pressure is 85
kPa and the wet bulb temperature is 25ºC in all three cases, but that the dry bulb
temperatures and consequently the gaps between the dry and wet bulb temperatures
increase. Firstly, it is noticeable that the heat content is exactly the same in each case, and
this bears out the statement in Note 11 that the sigma heat of air depends only on the wet
bulb temperature and the barometric pressure and is not affected by the dry bulb
temperature. It is also noticed that, as the gap between the dry and wet bulb temperatures
increases, the moisture content and the relative humidity decrease, but the specific volume
increases because of the higher dry bulb temperature.

        (c) Comparing line 1, 7 and 18 in which the air is saturated at 15ºC in each case but
the barometric pressure increases from 85 kPa to 100 kPa and 115 kPa, it is seen that the
heat content, moisture content and specific volume all decrease progressively. This means
that air at the same temperatures but at a higher pressure contains less heat and less water
vapour but is denser.

        (d) Comparing line 7 with 8, 9 with 10 and 11 with 12, we see that in all cases the
barometric pressure is 100 kPa and the air is saturated, but a kilogram of air at 16ºC
contains 2.5 kilojoules more heat that at 15ºC, at 26ºC, 3.9 kilojoules more than at 25ºC, and
at 38ºC, 6.9 kilojoules more than at 37ºC. Thus, it does not always require the same amount
of heat to raise the temperature of one kilogram of air through 1ºC. The reason for this is
that different amounts of water are associated with the kilogram of air at different

        For rough mental calculations, it is useful to remember that it requires 4 kilojoules to
heat one kilogram of dry air and the water vapour associated with it by 1.C wet bulb when it
is at about 26ºC wet bulb and 100 kPa pressure.

        (e) Comparing line 18 with 20, 21 with 23 and 24 with 26, we see that in all case, at
constant pressure, the wet bulb temperature is raised by 1ºC while the moisture content
remains the same. In the first case the dry bulb temperature increases by 2.6ºC, in the
second case by 3.7ºC, and in the third case by 5.9ºC. This is the phenomenon which is
observed whenever air is heated without the addition of moisture, e.g. when passing through
a hoist chamber or an electric motor. The dry bulb temperature always increases several
times as much as the wet bulb temperature, and the relative humidity decreases.

        (f) Lastly we will examine what happens when air travels down a perfectly dry shaft
by comparing line 7 with 15, 9 with 16 and 11 with 17. In each case the barometric pressure
is increases by 2.5 kPa, which is approximately what happens when air travels 210 metres
vertically down a shaft. Also in each case the moisture content remains unchanged, which is
what would happen in a perfectly dry draft. The air, in travelling down the shaft, loses some
of its potential energy, which is converted into heat. One kilogram of air moving 210 m
vertically down a shaft loses 2060 Joules of energy. (2060 J = 1kg × 210m × 9.8 m/s 2).
Thus, the heat content of a kilogram of air is increased by 2.060 kJ. This amount of heat was
added in lines 15, 16 and 17. It was then found that in each case the dry bulb temperature
increased by 2.1ºC, (i.e. bout 1ºC per 100m) while the wet bulb temperature increased by
amounts varying from 1.1ºC to 0.7ºC, depending on the initial temperature.

       When the various sources of heat affecting the ventilation air were listed at the
beginning of Note 7, auto compression was mentioned first because it is one of the most

important sources in deep mines and because it is unavoidable. Even if there is no flow of
heat from the rock or any other sources. The wet bulb temperature increases by about 1ºC
with every 300 m increase in vertical depth. In a mine 3000 m deep this increase is very
serious, especially in summer when the air leaves surface at a wet bulb temperature of well
over 15ºC.

       This auto compression is also sometimes called adiabatic compression. The term
"adiabatic" means "without the addition or removal of heat from external sources".



      Psychrometry has many practical applications in mine ventilation. In this note a few
examples will be worked in order to indicate some of the uses of this science.

       Example 10: Temperatures are measured at an underground station in a deep mine
and at 1000 m intervals along an apparently very dry haulage carrying air from the station.
The barometric pressure on this level is 100 kPa.

                                            TABLE 6

                   Observations                         From Psychrometric Charts

        Distance from        Temperatures          Moisture Content       Sigma Heat
           Station             Wet/Dry                   g/kg                kJ/kg
              m                   ºC
               0               24.0/32.0                 15.7                70.9
             1000              25.0/32,0                 17.4                75.0
             2000              25.5/34.0                 17.4                76.8
             3000              26.5/38.0                 17.4                81.0

        A study of the temperatures alone would only indicate that more heat is picked up by
the air in the first and the last sections than in the middle section. The psychrometric
analysis, however, shows that each kilogram of air picks up 1.7 grams of moisture in the
first section and nothing in the other two sections. Thus, even though the first section
appears perfectly dry, this is the case only because all the moisture, which seeps from the
rock, is immediately evaporated. Slightly more heat is actually added to the air in the last
section than in the first section – 4.2 kJ as against 4.1 kJ. This may be because this section
has not had sufficient time to be cooled, or because there may be less air is flowing in it, but
further water control measures would be wasted in this section.

        Now what improvement could be expected if the first section could be effectively
dried out? The humidity of the air would be then remaining 15.7 grams per kilogram
throughout. If it is assumed that the rock will not be cooled any further, then the dry bulb
temperature at 3000 m, will probably remain 38ºC. Reference to the psychrometric charts
shows that the wet bulb temperature at this point will then be only 25.5ºC or 1ºC less than

       Example 11: 10 m3/s at 20/27ºC and 100 kPa enters a hoist chamber in which a
3000 kW motor is running at 95 per cent efficiency. What will the temperature of the air
leaving the chamber be?

       Of the 3000 kW, 95 per cent is used efficiently and remaining 5 per cent (i.e. 150
kW) is converted into heat. Thus the amount of heat produced is 150 kJ/s.

        Now, from the psychrometric charts it is found that air at 20/27ºC and 100 kPa has a
sigma heat of 56.4 kJ, a moisture content of 11.9 g and a specific volume of 0.878 m 3, all
per kilogram of dry air.

       Thus 10 m3/s of air is equivalent to 10/0.878 = 11.4 kg of the dry air per second.

       Each kilogram of dry air must before remove 150/11.4 = 13.16 kJ. A kilogram of the
entering air contains 56,4 kJ and the leaving air must thus contain 56.4 + 13.2 = 69.6 kJ per
kilogram. In a dry hoist chamber no water is added to the air, and its humidity is thus still
11.9 grams per kilogram.

       From the psychrometric chart for 100 kPa it can now be found that air containing
69.6 kJ and 11.9 grams of vapour per kilogram has a temperature of 23.8/40.5ºC.

       Example 12: 30 m3/s of saturated air at 25/25ºC and at 100 kPa is passed through a
cooling plant and is cooled to 15/15ºC. How much heat and water is removed?

   The psychrometric values for these conditions all appear in lines 7 and 9 of Table 5 in
Note 12.

       The mass flow of dry air entering the plant is 30/0.885 =33.9 kg/s. The amount of
heat removed from each kg of air is 74.9 – 41.6 = 33.3 kJ. Thus the total rate of heat
removal is 33.9 × 33.3 =1130 kJ per second or 1130 kW. The air enters with 20.3 –10.8 =
9.5 grams per kg, or about half the water vapour, is removed. This amounts to a total of 33.9
× 9.5 = 322 grams per second or 0.322 kilograms of water per second or 0.322 litres of
water per second

        This is quite an appreciable a stream of water, which will flow from the cooling
plant and which has been derived from condensation or artificial rain-making.

        Example 13: During a cold winter morning 500 m3/s of air enters the collar of a
large surface shaft at 4/4ºC and a barometric pressure of 85 kPa. At the bottom of the
vertical shaft, 2000 m deep, the temperatures are 17/27ºC at the pressure is 105 kPa. After
passing down a sub-vertical shaft, along intake haulages and up some lines of deep level
stopes, the air reaches a point at the same elevation where it is at 24/24ºC and 85 kPa. How
much heat and water is absorbed in each section of the circuit?

      This problem combines several of the problems previously discussed and
consequently the answers are given below without showing all the calculations.

                                       TABLE 7

                                        Top of               Bottom of          Above           Top of
                                        Downcast             Downcast          Stopes           Upcast
  Temperature ºC                         4/4                 17/27             32/32             24/24
  Barometric pressure kPa                85                   105               105               85
  Total Air Volume m3/s
  (mass flow rate specific volume)      500                   440               463              5 51
  Specific Volume m3/kg                 0.946                0.831             0.875             1.041
  Mass Flow Rate of Dry Air kg/s        529                   529              529               529
  Moisture Content g/kg                 6.1                     7.7             29,6             22.7
  Difference g/kg                                    +1.6             +21.9             -6.9
  Total Moisture differences l/s                     +0.85            +11.58            -3.65
  Sigma Heat kJ/kg                      19.2                  46.0             103.7             79.8
  Difference kJ/kg                                   +26.8            +57.7             -23.9

       (a) It is seen that the volume of air changes considerably between one position and
another, but the total mass of dry air remains exactly the same through. However, the mass
of water vapour contained by the air does change.

        (b) The air evaporates 0.85 litres of water per second in the downcast shaft and
11,58 litres per second in the lower airways and in the stopes. However, 3.65 litres per
second condense out of the air while it rises to surface in the upcast shaft. This latter water
in liquid form considerably increases the work of the fan, which has to force the air out of
the upcast shaft, as will be seen later.

       A net quantity of 8.78 litres of water per second is thus removed from the mine by
the ventilation current - more than is pumped out by some dry mines.

        (c) In the downcast shaft 26.8 kJ are added to each kilogram of air. Of this 2000 ×
98 × 0.001 = 19.6 kJ are due to auto-compression (see Note 12) and only 7.2 kJ come from
the rock or other sources such as pump columns and compressed air columns.

        (d) In the upcast shaft the heat content of a kilogram of air is reduced by 23.9 kJ. Of
this 19.6 kJ are again due to auto compression, but the remaining 4.3 kJ are lost because of
the heat flow from the air to the rock, because in this shaft the air is hotter than the rock.

       (e) The total amount of heat removed from the mine is 529 × (79.8 –19.2) = 32000
kJ/s. One kilo joule per second is equivalent to one kilowatt (see Note 7). Thus the heat
energy removed from this mine by the ventilating air is equivalent to 32000 kilowatts = 32

        Although this last example has been rather simplified by omitting the effects of fans,
it gives a fair picture what can happen in any deep mine.


                                  CHAPTER 3 (MINE DUST)

3-1 DUST (NOTE 14)

                       Note 6 started off by stating that air is required in a mine for
       four reasons- to supply oxygen for breathing, to remove heat and to dilute
       and remove dust and gases. Having discussed the first two items, we can
       now turn our attention to the third.

         Dust is formed by reducing materials to small size. Processes like grinding,
crushing, blasting and drilling produce dust particles of sizes varying from very fine to very
coarse. The size of the dust particles formed from any particular material depends on the
force applied per unit area of the material. Thus a blow with a large hammer on a rock will
split it into large pieces and will create mainly coarse dust, while a blow of the same
strength on to a sharp chisel will break a small bit of the same rock into fine dust particles
because the same force is directed onto a much smaller area.

         In nature dust is caused mainly by grinding, such as the grinding action of
windblown dust on rocks and of waterborne rocks on each other as in rivers and on the
ocean shores. In all these cases the force of impact is rather limited and consequently most
of the natural dusts are fairly coarse as can be observed in beach, river and desert sands.
Nature protects the lungs of humans and other animals against this natural dust by supplying
pre-filters, which are very effective on coarse dusts. These filters consist of hairs in the nose
and constant moistness of the linings of the nose, mouth and windpipes. Furthermore these
filters are self-cleaning. The windpipes are lined with myriads of cells with whip like
appendages, called cilia, which carry upward any foreign bodies that chance to touch the
wet mucus-bathed linings. All the mucus is always moving towards the nose and mouth.
Thus the dust eventually reaches the mouth from which it is either expectorated or
swallowed. The inhalation to really coarse dust causes coughing and sneezing, which
expedite its removal.

        In industry, dust is caused by much more powerful and concentrated forces such as
blasting, drilling, grinding, scraping, crushing and milling. The result is that much finer
dusts are formed for which nature never catered in the design of its pre-filters. Some of this
very fine dust consequently does manage to reach the finer passages of the lungs and to
remain there. Any such accumulation of dust in the lungs causes an unhealthy condition
called pneumoconiosis. This word is derived from two Greek words meaning "lung" and
"dust". (The words pneumonia and konimeter were also derived from the same Greek
roots.) Various kinds of pneumoconiosis are caused by different kinds of dust, and some are
more severe than others. Anthracosis is caused by breathing coal dust; asbestosis by
asbestos dust; and silicosis by silica dust.

       Silica or silicon dioxide (SiO2) is most commonly known to the layman as quartz but
also occurs in other forms, e.g. tridymite and cristobalite. Quartz itself is given various
names such as rock crystal, amethyst, smoky quartz, milky quartz, etc., depending on its
colour, which is caused by various impurities, and its shape. Silica and silicates (silica
combined with some other base) occur very freely in nature and actually form a very high

percentage of the crust of the earth. The gold-bearing ore in South African mines contains
up to eighty per cent silica.
         An accumulation of only a few grams of silica dust is normally found in the lungs of
miners who died after contracting an advanced stage of silicosis. One definition of silicosis
is that it is "a disease due to breathing air containing silica, characterised anatomically by
generalised fibrotic changes and the development of miliary nodulation in both lungs, and
clinically by shortness of breath, decreased chest expansion, lessened capacity for work,
absence of fever and increased susceptibility to tuberculosis".

       The smallest dust particle that can be seen with the naked eye under ideal
conditions- e.g. a black particle on a sheet of white paper with good lighting-is about 25
microns. With a good light microscope one can see down to just under a quarter micron.

        Natural dusts as seen in dust storms and on gravel roads with cars passing are
coarse, consisting predominantly of particles larger than 40 microns. Mine dusts, however,
are very much finer and, except under extremely bad conditions which normally last only
for very short periods, are completely invisible. It has been found that nearly 80 per cent of
the dust particles in mine air are smaller than 1 micron, nearly 20 percent are between 1 and
4 microns and only 4 percent are larger than 4 microns. The dust collected from the lungs of
deceased silicosis has a similar size distribution, except that there are rather more of the
finer and fewer of the larger particles.

        In the fight against silicosis the primary aim is to prevent dust becoming airborne.
As a second line of attack, that dust, which cannot be kept out of the air, is either exhausted
directly at its source, or filtered, or it is diluted as effectively as possible by means of
ventilating air.



               When a car travels fast over a gravel road it raises a thick cloud of
       dust, but after ten seconds or so the cloud has usually disappeared. The
       reason is that most of the dust is coarse and rapidly falls to the ground
       where it remains settled until disturbed by another car or a very strong wind.

        All free falling bodies, even the smallest, obey the laws of gravity. If a small stone
and a large stone or a stone and piece of lead are dropped over the edge of a balcony at the
same time, they can be heard to strike the ground at the same time. If, however, a sheet of
paper is released at the same time as a stone it will wave backwards and forwards and reach
the ground a long time after the stone.

       If the paper is now torn in half and one half rolled into a ball, the ball will reach the
ground much sooner than the other half sheet if the two are released simultaneously. From
these experiments it can be concluded that the rate of fall of a body is not dependent on its

mass or its density but that it is affected by its shape when falling through air, because, in
falling, a flat body has to push aside a greater volume of air than a spherical one. Bodies
falling through air are thus not free falling bodies, but in a vacuum space it would be found
that a feather or a one micron dust particle would fall just as fast as a piece of lead.

     Accurate observations on a freely falling body show that it falls:

     4.9 metres in one second (4.9 × 12)
     19.6 metres in two second (4.9 × 22)
     44.1 metres in three second (4.9 × 32)
     78.4 metres in four second (4.9 × 42)

       It will be seen that these figures conform to the quadratic equation s = at2 in which 's'
represents the total distance, 'a' the distance travelled in the first second and 't' the time in

        From these figures it can also be calculated that the speed at which the body is
travelling is:

        9.8 metres per second at the end of the first second,
        19.6 metres per second at the end of the second second,
        29.4 metres per second at the end of the third second,
        39.2 metres per second at the end of the fourth second,

         After each second the speed is thus 9.8 metres per second greater than after the
previous second. The increase of speed with time is called acceleration. Speed is distance
divided by time and acceleration is the increase in speed in unit time. We can therefore say
that the acceleration of a freely falling body due to gravitational forces is 9.8 metres per
second, per second. This is a very important figure in physical science and is denoted by the
letter 'g'.

        The value of 'g' is exactly twice that of 'a' which was used in the equation       s=
 2                                                              2
at , and therefore the equation can now be written s = ½ gt , and the speed or velocity 'v'
after any required time or distance of fall can be calculated from the formulae v = gt and v =
√2gs .

        Using this last formula it will be seen that if a raindrop or a hailstone were to
fall freely from a cloud 1500 metres above the earth, it would reach a speed of 172
metres per second or 618 kilometres per hour, which would probably be sufficient
to kill any living being. The same would apply to water or stones falling down a
1500 metre deep mine shaft. However, in both cases, the free fall of the bodies is
hampered by the resistance offered by the air. The air resistance depends on the
density of the air and on the shape, size and speed of the falling body.

       For any set of conditions a falling body will after a time reach a maximum
speed at which the gravitational pull and the air resistance are in equilibrium. This
speed is called the terminal settling velocity of such a body. The terminal velocity
is greatest for a body whose mass is largest compared with its surface offering
resistance and therefore large bodies will reach a greater speed than small bodies
of the same shape and density. Furthermore the falling of very small bodies is
affected by other factors such as Brownian Motion (due to the buffeting action of
the gas molecules) convection currents, non-uniform temperatures and other
disturbing forces.

        Taking all these factors into account, the following terminal velocities have been
calculated for quartz:

                               TABLE 8

          Particle Size                Terminal Velocity
            Microns                  m/s              Km/h
        500000               178                           640
        25000                40                            140
        2500                 12.7                           46
        1000                 6.0                            22
        250                  1.5                             5
        100                  0.4                             -
        50                   0.13                            -
        20                   0.02                            -
        10                   0.005                           -
        5                    0.0015                          -
        1                    0.00005                         -

       A one micron dust particle liberated 2 metres above the footwall will thus take about
10 hours to settle on the floor, while plainly visible 100 micron road dust would settle in
about 5 seconds.

        Closely connected with the concept of terminal velocities is that of conveying
velocities. When an exhaust system is designed to draw dust or waste products such as
wood chips away from the point of production, or to convey material such as grain, the air
velocity in the pipes must at least equal the terminal settling velocity of the largest particles
of the material that has to be removed. Usually these velocities are of the order of 10 to 20
m/s for exhausting dust and up to 35 m/s for conveying heavy loads of materials such as
sand and cement.

        The best visual demonstration of the different conveying velocities required for
different sizes of particle is sometimes encountered in an upcast shaft where small drops of
water can be seen travelling upwards with the air at different speeds, while large drops can

also be seen falling steeply to the floor while smaller ones follow a flat parabolic trajectory
to the floor and still smaller ones continue flowing horizontally into the fan. Modern surface
exhaust fan drifts are often constructed with a wide portion so as to reduce the air velocity
and thereby allow much of the water to fall out before the air enters the fan.



        One cubic millimetre of rock crushed to cubes of one-micron size would yield 1000
million dust particles. In drilling a 40 mm hole 1 metre deep, the volume of rock removed is
1.2 million cubic millimetres; drilling such a hole takes less than 10 minutes. Assuming an
airflow in the stope of 20m3/s, such a hole could add 100,000 particles of dust per
millilitre (p/ml) of air. Fortunately, however, very much less than one percent of the rock
drillings are as small as 1 micron in size and most of then are prevented from becoming

       The largest numbers of fine dust particles are produced in a mine by blasting;
additional dust is produced by mechanical actions such as drilling, scraping, barring,
lashing, tipping and loading.

        Some of the fine dust produced during blasting is carried away by the air steam,
which is not breathed by anybody, but a large amount is trapped with the broken rock. Some
of the coarse particles, which have become airborne, settle out on the footwall and some of
the finer particles collide with each other and coagulate to form larger particles, which then
settle out.

        Nearly all of the very large amount of dust produced by drilling is caught by the
water flowing down the drill steel and comes out of the blasting hole as sludge.
Unfortunately all the dust is not caught in this way because some compressed air usually
leaks past the piston and finds its way down the drill steel to the bottom of the hole where it
collects some dust before escaping to the atmosphere. The modern sealed-spline rock drills
allow less air to escape than the old exposed spline machines. The front head release ports
allow some of the air, which gets past the piston to escape without passing down the hole.

       The dust created by scraping, barring, lashing and loading can to a large extent be
kept out of the air stream by ensuring that the rock is kept sufficiently wet.

        Thus far only the creation of dust has been mentioned and we have seen that most of
it does not become airborne but for various reasons remains on the footwall. However if this
dust is dry or allowed to become dry it can easily be made airborne by disturbing it
mechanically (scraping, lashing, sweeping, walking, tipping) or by blowing air over it at
high velocities. Dust can also be caused by evaporating water, which contains dust. In this
latter connection it bears mentioning that clean water contains anything up to ten million
particles of dust per millilitre and that dirty water can contain ten or twenty times as much.

Thus the evaporation of water from a fogging rock-drill or a tip atomizer can increase the
dust content of the air.

        In all cases where rock is to be handled or is likely to be disturbed it is essential that
it should be damp, but care must always be taken not to waste water because, apart from the
cost of the wastage, this may be the indirect cause of more dust becoming airborne. Apart
from the minimum amount of about 0.05 litre/s, which should be used by each rock-drill,
and water issuing from fissures, nearly all running water observed in a mine can be regarded
as wastage.
        If sufficient water is applied to the broken rock by means of a spray to wet all of it,
but without allowing any water to run to waste, very little dust will become airborne as a
result of lashing and scraping. However, if the full bore of an open hose is turned on a pile
of rock, some of the water does usefully wet the rock, but most of it finds its way through
the loose rock and runs down the foot wall and pours out of the box hole below the stope as
dirty water. It then flows along the drain, which, due to wastage at many other places,
periodically becomes dirty and over filled. When it overflows it deposits fine dust on the
footwall. After the drain has been cleaned and the footwall has dried off, this dust will again
be disturbed by people walking over it and by trains passing over it.

        Some of the water from the box-hole enters the cars, which are filled with rock.
Some of its drips out along the haulage and dust is made airborne as before. Some of it is
tipped into the shaft bin from which it enters the skips. Dripping from the skips again brings
fresh fine dust onto the shaft supports and eventually into the air stream. In this round about
way some of the dust created by drilling and blasting in a stope is thus liberated into the air
stream in the down cast shaft.- all because of the excessive use of water , which was
originally intended to allay dust.

        Once fine dust has settled it is not very easily disturbed by a steady wind velocity.
The footwall and sides of fan-drifts and even the protection screens on fans are often seen to
be thickly coated with dust. However, when a fan is stopped and restarted a cloud of dust is
often created for a short period while the speed of the air is increasing. This phenomenon
can be compared to what happens on an inclined conveyor belt. When the belt is in steady
motion its load of rock moves with it without any sign of slipping, but when it is stopped
and restarted there is often evidence that the load wants to stay behind or even slide down.

        Now, when a piece of rock covered with fine dust is allow to fall, as happen when it
is transferred from one conveyor belt to another or when it is dropped into a tip. It is subject
to gravitational acceleration and subsequently to a sudden stop or deceleration. These
processes release dust into the air and if the tip is upcasting this dust is then brought out into
the ventilation current. But even if the tip is not normally upcasting, the falling stones draw
air with them and thus induce a local downward air stream, which must again find its way
out of the tip and therefore causes an up draught in some other part of the tip. There are
three ways of reducing the amount of dust which is liberated into the ventilating current by
this process- making the tip opening as small as possible, reducing the vertical drop of the
rock to a minimum and lastly, but most expensive, drawing off and filtering air from
underneath the tip in order to counteract the up draught.


(Note 17 to 23 are omitted)

                         CHAPTER 4 (BASIC VENTILATION)


       Having discussed the properties of air and the reasons why a continuous supply of
fresh air has to be circulated through a mine, it is now necessary to discuss ways of
achieving this air circulation.

       It is perhaps difficult to imagine that ordinary still exerts any pressure at all because
we are surrounded by air and yet we do not feel as though our bodies are being compressed
to any extent. Normal air pressure at sea level is equivalent to that caused by mass of more
than 10000 kg resting on an area of one square metre. Thus every person carries a mass of
between 500 and 1000 kg of air on the horizontal portions of his head and shoulders! In
addition to this, a man standing with his feet on the floor of a swimming pool where it is 1.5
m deep would have at least another 20 kg of water resting on the top of each of his feet -
and he would bear it with a smile!

        The reason for these phenomena is that the static pressure of gases and liquids of
which the human body consists - can stand very great pressures if they are applied evenly
from all directions. If a pressure 200 times as great as sea level atmospheric pressure (i.e.
20000 kPa) is applied to water, the volume of water is reduced by only 1 per cent. Thus our
bodies do not feel normal atmospheric pressure and when the pressure on our bodies is
changed, for example when we go down a mine or dive under 2 or 3 m of water, we feel no
effect except for a temporary feeling of pressure on the eardrums which disappears as soon
as the pressure inside the eardrum has been adjusted.

        When one sucks air out of a balloon, the balloon collapses and it is virtually
impossible to part the two sides of the flattened balloon because the air pressure is now
acting only on the outside. Similarly, if one sucks air out of a bottle, which is too strong to
collapse, one's tongue becomes tightly drawn to the opening of the bottle because there is
more air pressure acting on its one side than on the other. If all the air were withdrawn from
the bottle and the opening had an area of 1 cm2; the pull would be sufficient to suspend a
bottle weighing more than 1 kg from one's tongue.

        Atmospheric pressure cannot be measured by means of a scale or a spring balance,
because the pressures on all sides balance not. The simplest way of measuring the pressure
is invented by the Italian scientist Torricelli more than three centuries ago.

        If a glass tube which is closed at one end is filled with mercury and the open end is
closed with a finger and placed under some mercury in a vessel and then released, the
mercury in the tube does not fall to the level of that in the vessel, but stands at a height of
about 760 mm above it (at sea level). The upper end of the tube has no air in it and
therefore no air pressure acting on the mercury, while the atmosphere presses on the

mercury in the open vessel and holds it up the tube at such a height that the pressure exerted
by the mass of the column of mercury is equal to the atmospheric pressure. (It must be
noted that it is the pressures- i.e. forces per unit area -which are in balance and not the
actual masses of air and mercury. Thus the relative sizes of the vessel and of the tube have
no effect on the height).

                                          Mercury       Vacuum

                      FIG. 2 Mercury Barometer

       Mercury is 13.6 times heavier than water. Thus 760 mm Hg = 10.34 m of water.
With a long enough tube it would thus be possible to use water instead of mercury in
Torricelli's barometer. However, this length of tube would make it unwieldy and therefore
mercury is always used in wet barometers. There are several types of mercury barometers
in which various ideas have been incorporated to ensure accurate measurement. In all
precise work temperature corrections are also made to compensate for changes in the
density of the mercury and for expansion of the glass tube.

        While mercury barometers are the most accurate and reliable and are therefore used
in laboratory work, they are not easily portable and are consequently not suitable for filed
work in mines. An Aneroid Barometer is much more convenient for this purpose. The
word aneroid is derived from two Greek words, “a” which means not and "neros" which
means wet. Thus an aneroid barometer is "not wet", which simply means that it contains no
liquid such as mercury. It consists essentially of an evacuated cylindrical box with elastic
faces, which are kept apart by a strong spring. The slightest movement of the top of the box
caused by changes in atmospheric pressure is transmitted to a pointer on a scale from which
the pressure can be read directly after calibration against a mercury barometer.

        Various types of a very accurate aneroid barometer have been evolved. They are
used for all kinds of work, such as determining the altitude of an aeroplane or of a
mountain; in mine ventilation work they are used for measuring pressure losses in the air

        So far only absolute static pressures have been discussed, i.e. the pressure due to
the total weight of an atmosphere of still air. Very often in mine ventilation work it is more
important to know the difference between the static pressures at two points - the so-called
gauge pressure. This is easily measured by means of a U-tube (manometer) containing a
liquid and having its two limbs connected to the higher pressure (positive side) and will rise

on the other side (negative side). In such cases we walk of positive and negative static
pressures as a matter of convenience, but the absolute static pressure is always positive.

       Although pressures and pressure differences are thus conveniently measured by
meaning the height of a column of liquid, it must be clearly understood that the unit of
pressure is the Pascal. The pressure difference measured by means of a manometer:


        where p = pressure in Pascal's
              g = 9.8 m/s2
             w = density of liquid, kg/m3 (1000 kg/m3 for water)
             H = difference in liquid levels, m

       Thus for a difference of 1 mm (0.001 m) of water,
          P = 9.8 ×1000 × 0.001 Pa
             = 9.8 Pa

        Thus when using a water manometer in mine ventilation work, it is very nearly true
to say that each millimetre of difference in the water levels represents a pressure difference
of 10 Pascals.



       As explained in the previous Note, we do not feel the pressure of still air. However,
we do feel the pressure of moving air because it acts only in one direction. A strong wind
not only ruffles our hair and clothes, but it is also difficult to walk against. Very strong
winds of gale force can blow people off their feet, blow roofs off houses and topple giant

        Moving air thus exerts a considerable pressure, and this pressure is appropriately
called velocity pressure. When travelling in a car at 40, 80 and 120 km/h one can actually
feel the strength of these pressures by sticking out one's arm and holding one's hand with the
palm facing the direction of travel. One soon realises that doubling the speed of travel not
only doubles the velocity pressure, but increases it considerably more - actually four times,
while trebling the speed increases the pressure nine times, etc. Thus the velocity pressure
increases as the square of the velocity of the air.

        Static pressure represents the potential energy of the air, while velocity pressure
represents the kinetic energy of the air. The algebraic sum of the two is the total pressure
of the air. Velocity pressure always has a positive value while a static pressure difference
relative to other air can be either positive or negative. To obtain total pressure, a positive
static pressure thus has to be added to the velocity pressure, while a negative static pressure
has to be subtracted from velocity pressure.

        When a tube connected to one limb of a U-tube containing water is held in such a
position that the movement of the air has no effect on it, and then only the static pressure of
the air is acting on it. However, if it is held in such a way as to face directly into the air
stream, both the static pressure and the velocity pressure of the air act on it and thus the total
pressure is measured. The velocity pressure is determined by subtracting the static pressure
from the total pressure, or by measuring it directly by means of a Pitot tube.



       FIG. 3 Pitot tube and manometer (U – tube)

        A Pitot tube consists of two concentric tubes of which the inner has an open end A
(see Figure 3) which is held so as to face directly into the air-stream, while the outer tube is
sealed at the end and rounded so as not to cause turbulence, but has several small holes
drilled into its side at B. Because the air flows smoothly past these holes, the velocity
pressure has no effect on than and only the static pressure is measured. If the two tubes are
connected to the two limbs of a manometer, the pressure indicated by the difference in
water levels, C is equivalent to the velocity pressure.

        It can be proved by means of basic scientific theory that, in addition to being
proportional to the square of the velocity, the velocity pressure is also directly proportional
to the density of the air. From this theory the following two useful formulae are obtained:

       VP = V2w/2
       V = (2VP/w)1/2

   Where in each case:

       V = air velocity (m/s)

   VP= velocity pressure (Pa)
   w = air density (kg/m3)

        Using these two formulae it is easy to calculate the velocity corresponding to any
velocity pressure, and vice versa, if the air density is known.

        In practical mine ventilation work it is often convenient to do rough mental
calculation on the spot and it is then useful to remember that air velocity is approximately
40 m/s when the velocity pressure is 1 Pa. (This is exactly true only when the air density is
1.25 kg/m3.)

        Because the velocity pressure is proportional to the square of the velocity, it is then
easy to estimate other relationships as follows:

                                      TABLE 16

           Velocity (m/s)                      Velocity Pressure (Pa)
               40 (144 km / h)                                          1000

                      30              (30/40)2 × 1000 = 9/16 × 1000 = 560

                      20              (20/40)2 × 1000 = 1/4 × 1000 = 250

                      10              (10/40)2 × 1000 = 1/16 × 1000 =      60

        The Pitot tube is the basic instrument for measuring air velocities. If it is properly
designed no corrections are required and it can be used directly for calibrating other
instruments such as orifice plate meters and venturi meters which depend on measuring the
static pressure difference between two points where the air is travelling at different
velocities, cup and vene anemometers which integrate the rotations of a wheel over a
measured time, and tracer gas techniques, and also indirectly for calibrating low velocity
instruments such as "swinging gate" anemometers and velometers, hot-wire anemometers,
kata thermometers, torsion anemometers and smoke and flame methods.



        Just as water flows only from a higher to a lower elevation unless assisted, air flows
only from a higher to a lower total pressure unless assisted by a fan or some other appliance
or by a density difference.

        The flow of a fluid (liquid or gas) is in the first place retracted by its internal
resistance or viscosity, which is due to the cohesion and interaction between its molecules
and which sets up frictional resistance between layers of the fluid moving at different
velocities in relation to one another. Water flows much more freely than honey or syrup
because it is much less viscous. Some gases are also more viscous than others. While the
viscosity of liquid decreases with increasing temperature e.g. hot syrup flows more freely
than cold syrup, it is interesting to note that the viscosity of a gas increases with increasing

         When a fluid flows slowly its movement is laminar, which means that different
layers of the fluid flow smoothly over each other. At higher velocities the flow becomes
turbulent, which means that the molecules move at random, thereby causing eddies which
result in a greater dissipation of energy.

        Osborne Reynolds (1885) carried out a classic series of experiments in which he
released thin lines of colouring matter in different fluids flowing through fine tubes. He
found that, up to a certain velocity, the colouring matter remained in a distinct straight line
(laminar flow), but that when the velocity was increased to a certain figure, the so-called
critical velocity, the colouring matter broke up into writhing streaks and the whole fluid
stream became coloured (turbulent flow).

        By forcing liquid through a very thin tube, Reynolds also determined that the
required pressure was directly proportional to the velocity when below the critical velocity,
but proportional to the square of the velocity when above the critical velocity. He also
found that the value of the critical velocity varied with variations in fluid density and
viscosity and the size of the tube or duct. He devised a coefficient, now known as the
Reynolds Number, by combining all these factors. It is a dimensionless coefficient, which is
obtained from the formula:

         Re = VDw / μ

   Where Re = Reynolds Number
         V = velocity (m/s)
         D = pipe diameter (m)
         w = density (kg/m3)
         μ = dynamic viscosity (Ns/m2)

        When the Reynolds Number is less than 2000 the flow is laminar, when it is more
than 4000 the flow is always turbulent and between these two values the flow can be one or
the other depending on the roughness of the pipe and on the upstream turbulent in the fluid.
In nearby all cases met in mine ventilation practice, airflow is fully turbulent. Possible
exceptions are the flow of low velocity air through the small apertures of a filter bag, air
leaking through packed ventilation walls and air drifting through worked out areas.

         Reynolds Numbers have a very important bearing on scale model work. When a
model is made of a projected shaft, for instance, in order to predict its resistance to airflow,
it is not only essential that it must be a true geometrical scale model in all respects, but also
that the test velocity must be such that the Reynolds number will be of the same order as
will pertain in the actual shaft.

       It has been started above that the Reynolds Number is a dimensionless coefficient.
It may be useful to explain at this stage what dimensions mean.

        Definitions of all units are based on metres, kilograms and seconds, the basic units
of length, mass and time.

   The dimension of distance is length = m
   The dimensions of area are (length × length) = m2

   The dimensions of volume are (length × length × length) = m3
   The dimensions of density are (mass/volume) = kg/m3 = kg.m-3
   The dimension of force is the Newton (N),
   Being equal to 1 kilogram metre per second squared = kg ms-2

   The dimensions of specific gravity are (mass/mass) = 1
   Specific gravity is thus dimensionless
   The dimensions of air flow are (volume/time) = m3s-1
   The dimensions of pressure are (force/area) = Nm-2
    (1 Pa = 1 Nm-2)
   The dimensions of velocity are (distance/time) = ms-1
   The dimensions of acceleration are (velocity/time) = ms-2
   The dimensions of dynamic viscosity are (force × time/area) = Nsm-2

   The dimensions of Reynolds Number are

      VDw/μ = velocity ×length × density

              = ms-1 m kgm-3
                 Ns m-2

              = kg ms-2

              = kg ms-2
                kg ms-2

        Reynolds Number thus has no dimensions - it is purely a number - and is therefore
called a dimensionless coefficient.



        When air flows through a duct, be it a ventilation pipe, a rock tunnel, a shaft with
steelwork or timber in it, or a stope, the pressure required to force or draw the air through
the duct depends not only on the internal air friction, but also on the size, length and shape
of the duct, the roughness of its sidewalls, the nature of the obstructions in it, and the
velocity and density of the air. Theoretically the Reynolds number is also important, but in
practical mine ventilation work this dependence is small since the flow is always fully

       All the above factors have been incorporated in the flow equation for air at standard
density that was proposed by Prof. J. J. Atkinson in 1854. This is the well-known
Atkinson Equation:

     p = KCLV2

   Where p = pressure loss (Pa)
        K = friction factor (Ns2/m4)
        C = perimeter (m)
        L = length (m)
        V = velocity of air (m/s)
        A = cross-sectional area (m2)

         These dimensions are rather complicated and confusing and are therefore seldom
used in practice. It is useful to try to memorize the few values given in Table 17. For
specific work an accurate factor must be determined by experiment or from a available
literature, but for rough calculations these figures can be used and intermediate values

                                           TABLE 17
                                 Friction Factor, K = (Ns2/m4)

                                   Airway                                K

             Ventilation piping                                      0.003
             Concrete linked empty shaft                             0.004
             Straight rock tunnel                                    0.01
             Concrete lined shaft with streamlined buntons           0.025
             Concrete lined shaft with R.S.J buntons                 0.05
             Heavily timbered rectangular shaft                      0.08

       Just as it is earlier for a lorry to transport a load of straw than a load of sand, it is
also easier to convey light air than to convey dense air along a duct. Actually the pressure
drop "p" is in direct proportion to the density of the air, and consequently a correction has to
be made to Atkinson's equation by writing it as follows:

     P = K CLV2 × w
           A      1.2

       Where "w" is the density of the air in kg/m3

       Air volume or Quantity (Q) = Air velocity (V) × Cross sectional area (A) i.e
                Q    = V ×A
               V = Q/A

       This expression can be used in the Atkinson equation, to give
           p = K CL/A (Q/A)2 × w/1.2

              = K CLQ2/A3 × (w/1.2) where Q is in m3/s

       The equation is more often used in this form than in its original form.

        Atkinson's equation has numerous applications in mine ventilation, a few of which
will be illustrated in the following examples.

       Example 14: What will the pressure drop be along a 4m × 3m rock tunnel, 1500 m
long, when 60 m3/s of air at a density of 1.35 kg/m3 is flowing along it?

                p = KCLQ2/A3 × w/1.2

                 = 0.01 × 14 × 1500 × 602 ×1.35       (In this case C = 4+3+4+3 = 14)
                            123 ×1.2
                 = 492 Pa

       Example 15: What will the pressure drop be if everything remains as in Example
14, except that the air volume is 70 m3/s?

                 p = 0.01 × 14 ×1500 × 702 ×1.35
                            123 ×1.2
                   = 670 Pa

        It will be noted that the figures are exactly the same as in the first example, except
that the air volume is different. Because the pressure drop "p" on the left of the equation is
proportional to Q2 on the right-hand side of the equation, it is possible to obtain the answer
such more quickly by simply stating that the pressure drop with 70 m3/s flowing (p70) is to
the pressure drop with 60 m3/s flowing (p60) in the same ratio as the square of the quantities,

   p70           = p60 × (70/60)2
                  = 492 × (7/6) 2
                  =492 × 49/36
                  = 670 Pa

        In the same way the pressure drop for each of several other air volumes can be
rapidly calculated for this particular airway. When these values of pressure drop are plotted
on a graph against the quantity, the graph will have the typical shape shown below, and
because the most important properties or characteristics of the airway can be off from this
curve, it is called the characteristic curve of the airway or duct. It is also sometimes called
the system resistance curve of the duct.

                   High resistance          Medium resistance
                   duct characteristic      duct characteristic

                                                              Low resistance
                                                              duct characteristic


                                     FIG. 4

       Example 16: What will the pressure drop be when 3 m3/s of air at 1.2 kg/m3 density
flows along 600 m of 400 mm diameter ventilation piping?

       The perimeter of a circular pipe is given by the formula D where  is equal to
3.1416 or approximately 22/7, and D is the diameter.

        Note that if D is in mm, the answer will be also be in mm unless D as 0.4 m. The
area of a circle is given by A= r2 or A=  D2/4 where r is the radius of the circle.

     P = KCLQ2/A3

       = (K DLQ2/1)  (43 / 3 D6 )

       = (K LQ243 )/ ( 2 D5 )       ………... (1)

       = (0.003  600  32  43 ) / ( 2  0.45)

       = 10260 Pa = 10.26 kPa

      Example 17: What will the pressure drop be if, under the same conditions as in
Example 16, a 560mm pipe were used instead of a 400 mm pipe?


       As can be seen from equation (1) above, the pressure drop is inversely proportional to
the fifth power of the diameter. Therefore the answer will be:

   p560 = p400  (400/560)5

       = 10.26  0.7145
       = 10.26  0.186
       = 1.908 kPa

         This example illustrates the very important fact that a small increase in diameter
results in a large decrease in pressure drop.

       The number of interesting problems that can be solved by means of Atkinson's
equation is well nigh unlimited.



       Apart from internal resistance to the flow of air which is caused by viscosity (Note
26), and frictional resistance caused by the rubbing of the air against the sidewalls of the
duct (Note 27), the flow of the air is also retarded by the presence of obstructions and bends,
and by variations in the cross-sectional area of the duct.

        Obstructions in air ducts retard the flow of air in two ways. Firstly, the obstruction
causes a reduction in the free cross-sectional area and consequently an increase in air
velocity, which, as shown by Atkinson's equation, causes an increase in frictional resistance.
But secondly, and usually more seriously, the flow pattern of the air is disrupted by the
presence of the obstruction and this causes eddy currents, which result in a loss of energy.
The extent of the disruption of the flow pattern is dependent not only upon the size of the
obstruction, but also on its shape.

                  FIG 5                                             FIG 6

                 FIG. 7                                              FIG. 8

       It is obvious from the above diagrams that a 250mm thick aerofoil section installed
perpendicular to the air stream will cause much less obstruction than a 250 mm square
timber in the same position. It is also obvious that a continuous 250 mm wall support along
an airway will cause less resistance than a number of 250 mm square timber supports in line
over the same distance.

        The old type of timbered rectangular shaft has a very high resistance to airflow (and
consequently a very high friction factor as shown in Note 27) because of three reasons - the
large percentage of the area obstructed by timbers, the bluff shape of the timbers and the
roughness of the sidewalls. It has often been found that an old timbered shaft has a lower
friction factor than a new one because the sharp corners of the timbers have been worn
away by falling spillage and in some cases because accumulated spillage has given it a
rounded upper profile.

        The more modern type of concrete-lined circular shaft with I --section steelwork
(Rolled Steel Joists, commonly referred to as R.S.J's.) is a big improvement on unlined
timbered shafts because it improved two of the three factors affecting resistance to airflow.
Less steelwork is required because steel is stronger than timber and consequently a smaller
percentage of the total cross-sectional area is obstructed and the shaft sets are spaced further
apart. Also the concrete lined sidewall is much smoother than bare rock. If it had not been
for these two factors, the circular shaft would have been only slightly better than the
rectangular shaft because the R.S.J. button which offers two bluff profiles to the impinging
air, both at its top and its bottom, causes more resistance to airflow than a square timber of
the same size. In recent years this problem has been overcome by using steel buntons of
streamlined shape instead of R.S.J.'s and consequently the friction factor of circular shafts
has been approximately halved.
        When air travels around a bend it loses some of its momentum -- or internal energy
-- due to its inability to change its direction smoothly. If the bend is badly shaped, eddies are
caused and these result in a further loss of energy. The air pressure lost in a bend depends
on the velocity of the air and is proportional to the velocity pressure. For a 90º bend it varies
from about 10 per cent of the velocity pressure in a very gradual bend (mean radius equal to
four times duct diameter) to more than 100 per cent for a very sharp bend.
        Energy losses also occur when the velocity of a current of air is changed by a
variation in the size of the duct, and these losses can be considerable when the change is
abrupt. The acceleration of air requires a certain amount of energy, but this is negligible
when compared with the losses caused by eddy currents. When air enters a pipe it is
accelerated from a stagnant condition to the velocity pertaining inside the pipe and the
amount of energy lost -- the so-called entry loss-- depends on the shape of the entrance, as
can be seen from Figure 9.

                                        Entry loss as a percentage of velocity pressure
                                         90% in circular pipes.
                  Straight Inlet        125% in rectangular pipes

                                          50% in circular pipes.
                                          70% in rectangular pipes.
                  Flanged Inlet

                                         5-25% depending on included angle and length

                    Inlet Cone.

                                             3-15% depending on the radius

                       Flared Inlet
                                              FIG . 9

     Very serious losses occur when the velocity of the air is suddenly reduced. This is
commonly called shock loss and is comparable in nature to the energy loss, which occurs

when a fast-moving car collides with the rear of a slow-moving car. In an abrupt
enlargement to more than twice the previous size of the duct the full velocity pressure is
lost, while in a gradual conical enlargement, such as a fan evase, the loss can be as little as
10 per cent.



        Air moves from one place to another only when a pressure difference exists between
the two places. The volume of air that will move in unit time depends on the magnitude of
the pressure difference and on the resistance offered by the duct to the flow of air.

        When a temporary pressure difference exists between two points, such as between
the inside and outside of an evacuated bottle, air will flow when the stopper of the bottle is
removed, but the flow will last only a very short while until the pressures are equalised.
When a continuous air current is required to ventilate a place, a permanent pressure
difference has to be created by a fan or some other means.

        Example 18: How much air at a density of 1.35 kg/m3 will flow through a 4m 
3m rock tunnel, 1500 m long, if a pressure difference of 1 kPa exists between the two ends
of the tunnel?

      p = KCLQ2            w (Atkinson's equation)
            A3              1.2

   1000    = 0.01  14  1500  Q2  1.35
               123  1.2                       (p is in Pa, not kPa)

               ( K= 0.01 -- See Table 17, Note 27)

   1000    = 0.1367 Q2

     Q2 = 1000

          = 7314

     Q    = 85.5 m3 / s

   Thus 85.5 m3/s will flow through the tunnel.

        Example 19 : How much air would flow through this same tunnel if the pressure
difference were only 492 Pa?


       The answer to this question can be found in the same lengthy any way as above.
However, it is not necessary to repeat the whole calculation when only one of the
parameters is altered. It can be seen from Atkinson's equation that the quantity squared is
proportional to the pressure drop. By taking the square root of the both, we find that the
quantity is proportional to the square root the pressure drop. Thus:

      Q2    p

   Q  √p

    New Quantity = √ New Pressure Drop
     Old Quantity     Old Pressure Drop

      New Quantity = Old Quantity  √ New Pressure Drop
                                       Old Pressure Drop

            = 85.5  √ 492
            = 60 m /s

       This agrees with the answer obtained in Example 14 of Note 27.

        Work is done in moving car against a resistance and energy must be supplied in
order to do this work. The next question is then, how much energy must be supplied?

       In Note 7 it was stated that power is the r ate of doing work. A joule is the amount
of work done when a force of 1 Newton moves its point of application by 1 metre.

        When a cubic metre of air passes through an opening of one square metre at such a
velocity that a pressure loss is one Newton per square metre (1 Pascal) it is obvious that the
amount of work done is equivalent to one Newton metre or one joule because the force is a
Newton and the air is moved a distance of one metre. If the cubic metre of air, however,
passed through a two square metre hole with the same pressure drop, the air would only
have to move a distance of 0.5 metre. The work done in each square metre of the hole is
now only 1  0.5 = 0.5 J but the total work is again 2  0.5 = 1 Nm = 1 J.

       This amount of work done is thus independent of the cross-sectional area through
which the air passes. It depends only on the volume of air, which is moved and on the
amount of pressure or force, which is required. Thus when 60 m3 of air is moved through a
tunnel of any size over any distance and the pressure drop is          1 N / m 2 (1Pa), the
amount of work done is
           60 m3  1N / m2 = 60Nm = 60 J.

       If the pressure drop is 2 N / m2 the work done is
          60 m3  2N / m2 = 120Nm = 120 J.

       If the pressure drop is 492 N / m2 then the work done is

          60 m3  492N / m2 = 29 520 Nm
                            = 29 520 J
                            = 29.5 kJ

        If 60 m3 of air is moved every second and the pressure drop is 492 Pa then work has
to be done at the rate of 29520 watts = 29.5 kW. Thus the power required to move 60 m3 /s
of air at a density of 1.35 kg/ m3 through a 4 m  3 m rock tunnel 1500 m long (Example
19) is 29.5 kW and this is called air power. (See Note 7 for definition of watt.)

         From the reasoning given above, a simplified can be derived for calculating air

   Wa = pQ

   Where Wa = air power (kW)
         Q = volume flow (m3/s)
          P = pressure (Pa)

   This is very important formula, which must be memorized.



       When air flows work is done, and for work to be done energy (power) is required.
This power can be supplied by a machine such as a fan or compressor or by other sources
such as heat or falling water or rock.

       When air movement is caused by heat (normally supplied by the rock) or falling
water (normally from a fissure in the rock) it is usually called natural ventilation.


                             FIG. 10

         In a mine having a shaft and an adit into a hillside as in Figure 10, air will travel
down the shaft during summer and up the shaft during winter because the temperature of the
air in the mine shaft stays comparatively constant throughout the year due to the consistency
of the rock temperature, while the temperature of the outside air changes with the seasons.
During summer, when the outside air is hot, it is too light to balance the cool column of air
in the shaft and consequently air moves down the shaft and consequently air enters the adit
and upcasts through the shaft.

   Falling Water

                                              FIG. 11

        When two vertical shafts are sunk from the same elevation and connected at the
bottom, the effect of the outside air temperature is cancelled out and it does not cause air to
flow. In the old days a flow of air was included in such a mine by either pouring water
down the one shaft (which was unpopular because the water had to be pumped or bailed out
of the mine) or by making a fire at the bottom of the other shaft. See Figure 11.

        Once the flow of air has been started by one of these methods, the rock in the
downcast shaft would be cooled more than in the upcast shaft and the air would continue
flowing because it would also gradually become cooler in the downcast. Consider such a
case as applied to a deep mine:

                                                                TABLE 18

                                                                   Downcast Shaft                         Upcast Shaft
                            Temperature, ºC
             Depth, m

                                                                               Density Kg/m3

                                                                                                                     Density Kg/m3
                                                Pressure, kPa
                            Virgin Rock



  Top of                0         18                 81            10/10                 0.99           22/22                0.95
  Middle      1000                27                 92            16/16                 1.10           27/27                1.05
  of shaft
  Bottom      2000                36               103             22/22                 1.21           32/32                1.15
  of shaft
  Mean                                                             16/16                 1.10           27/27                1.05

           For ease of calculation the air has been assumed to be saturated throughout the

        In each shaft the mass of an air column with a cross-sectional area of 1 m2 will be
given by the mean density of the air multiplied by the depth of the shaft, thus 2000  1.10
= 2200 kg in the downcast shaft and 2000  1.05 = 2100 kg in the upcast shaft. To obtain
the pressure exerted by these two columns of air the masses must be multiplied by g, the
gravitational acceleration. Thus the downcast air exerts a pressure of 2200  9.8 = 21560
N/m2 or 21560 Pa and the upcast air exerts a pressure of 2100  9.8 = 20580 Pa.

        The difference between these two pressures is 980 Pa and the available natural
ventilation pressure is thus 980 Pa. It is obvious that the same answer can be obtained more
expeditiously by multiplying the difference between the two masses

           (2200 - 2100 =100) by 9.8, thus
            100 kg/m2  9.8 m/s2 = 980 kg ms-2 / m2
                                 = 980 N/m2
                                 = 980 Pa (See Note 26)

        The above method of calculation involves a considerable amount of work. In
practical mine ventilation work it often happens that an approximate value of the natural
ventilation pressure is required quickly. In such cases the author has found the following
method of calculation, which can be done mentally, to be very satisfactory.

        It is only necessary to remember two facts. The first is that approximately 0.085 m
of motive column is equivalent to 1 Pa. This simply means that an sir column of standard
density (1.2 kg / m3) and about 85 mm high exerts a pressure of 1 Pa. Secondly, the density
of air at constant pressure is roughly proportional to the absolute dry bulb temperature, the
effect of humidity being relatively small.

       In the case of the example given above, the mental calculation would be done as

The mean absolute dry bulb temperature in the downcast shaft is 273 + 16 = 289 K.

The difference between the mean dry bulb temperature in the two shafts is 27 - 16 = 11ºC
       The motive column will therefore be approximately
            = 11  2000 m
            = 1  2000 m
            = 76 m

           The natural ventilation pressure is therefore approximately

               76 = 900 Pa

       This answer is very near to the correct figure calculated above, but in cases where
the one air column is more humid than the other, discrepancies will be much greater .

        The amount of air flowing as a result of natural ventilation pressure will depend on
the resistance of the circuit. In the above cases, if the shafts and airways are large, it is quite
possible that 150 m3/s will flow through the mine, which means that the equivalent of nearly
150 kW of power is supplied by natural sources.

    Wa = 150  980 =147 kW
    Drive 180 m  2.5 m  2.5 m

                                                                                          52 level
          6.5 m /s
                                        0.9 m3/s                     6.5 m3/s

    Barometer 108.0 kPa                                              Raise 60m  1.8 m  2.5 m
                                   33/34ºC         30/30ºC

                                                                                         53 level

                                        5.6 m3/s
                                        FIG. 12

        The author once encountered the case illustrated in Figure 12 in a deep mine.

       There was a vertical difference of 45 m between the levels and the virgin rock
temperature was 40ºC. The barometric pressure was 108 kPa.

         Considered purely as an air splitting problem, one would have expected much more
air to take the short straight route along 52 level than the long roundabout way down to 53
level and back, but here just the opposite was happening. Calculation shows that in this case
the N.V.P is only 8.8 Pa but that is just sufficient to create the flow of 5.6 m 3/s along the
longer route.

                                              7.5 m3/s

                                      26/32ºC                         27/27 ºC

            10 m3/s                           2.5 m3/s                  10 m3/s
                                        FIG. 13

       The opposite could also happen in the case of two cross-connected raises off a dry
haulage if water were added to the air at the top as illustrated in Figure.13.

        Another interesting case experienced by the author was that of an abandoned vertical
shaft which normally carried a downcast air volume of about 40 m3/s. This air served to
keep an inclined shaft fresh and then joined the general mine upcast through olds outcrop
shafts and workings.

        The downcast quantity was always less in summer than in winter, but during one dry
summer period it suddenly stopped altogether, resulting in the inclined shaft being filled
with smoke at blasting time. Two water hoses were used for spraying water down the shaft
from surface. This immediately had the effect of starting the flow of air again and within
half an hour the flow had increased to 40 m3/s. The water was then turned off and no further
trouble was experienced.

      In one old isolated inclined shaft there was a strong upcast in the ladder way
compartment, probably due to heat from a compressed air main, while the other
compartments were all down casting.

        A very common occurrence in hot mines is the flow of air in dead ends. Cold air
flows in along the footwall, gradually becomes warmer and flows out along the hanging.
Such a flow of air has been seen to persist for over a thousand metres. Air temperatures near
the hanging were about 3ºC higher than near the footwall. Because flow of this type often
occurs it is essential, when measuring small air quantities by means of the smoke method,
not to rely on observations in the centre of the drive only, but to make check observations
near the hanging and near the footwall.

        In a mine with wet downcast shafts and airways the fans get considerable assistance
from natural pressure, while in a very dry mine it is even possible that the natural ventilation
pressure may act against the fans and that the flow may consequently reverse when the fans
are stopped.

       The effect of air-cooling plants on natural ventilation pressures also depends on the
placing of these plants. As a result of these last two last two factors, mines with wet
downcast shafts or with surface cooling plants are much less affected by fan stoppages than
mines with dry shafts and underground cooling plants.

        A kilogram of water (or rock) falling down a shaft 2000 m deep, loses 2000  1 
9.8 = 19600 J of potential energy. If the water falls straight down, without hitting the
sidewall or supports, all this energy, except for the small amount absorbed when it hits the
shaft bottom, is transferred to the air through which it passes. If 2 litres of water fell down
this shaft per second, the energy imparted to the air would be of the order of 2  1  2000 
9.8 = 39.2 kW.

       Part of this energy would cause the air to downcast while the remainder would be
converted into heat.

        Rock, which is tipped into an ore-pass system or discharged from a conveyor, can
similarly cause considerable amounts of air to flow because of the energy imparted to it, and
in such cases it is usually advisable to reduce the distance of free fall of the rock to a
minimum in order to prevent currents of dust-laden air.


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