Mine Air Conditioning and Ventilation

MINE AIR CONDITIONING AND VENTILATION

Planning Principles and Definitions Air Cooling Techniques
Sources of Heat Entering Mines Selecting a Mine Cooling Method
Heat Exchangers Mine Air Heating
Mechanical Refrigeration Plants Mine Ventilation

In underground mines, excess humidity, high temperatures, inadequate oxygen, and excessive concentrations of dangerous or toxic gases can lower worker efficiency and productivity, and can cause illness and death. Air cooling and ventilation are needed in deep underground mines to minimize heat stress and remove contaminants. As mines become deeper, heat removal and ventilation problems become more difficult and costly to solve.

PLANNING PRINCIPLES AND DEFINITIONS

Planning for mine air conditioning and ventilation is a two-step process. First, requirements must be projected. Heat loads and ventilating airflows must be determined. Second, a system must be designed with appropriate resources to provide those requirements. Definitions specific to mine air conditioning and ventilation are given below.

• Heat Stress. Heat stress is a qualitative assessment of the work environment based on temperature, humidity, air velocity, and radiant energy. Many heat stress indices have been proposed over the years. See Chapter 8 of the 2001 ASHRAE Fundamentals Handbook for a thorough discussion. The most common indices in the mining industry are the effective temperature (Hartman, 1997), air cooling power (Howes, 1997), and the wet-bulb temperature. The following wet-bulb temperature ranges were derived from references and experience at several deep western U.S. metal mines:
twb  80°F Worker efficiency is 100%.
80° < twb  85°F Economic range for acclimatized workers.
85° < twb  91°F Safety factor range; corrective action required.
91° < twb Only short-duration work with adequate breaks.
• Heat Strain. Heat strain is the human physiological response to heat stress. Effects include sweating, increased heart rate, fatigue, cramps, and progressively worsening illness up to heat stroke. Individuals have differing tolerance levels for heat.
• Reject Temperature. This important number, based on the heat stress/strain relationship, is the wet-bulb temperature at which air should be rejected to exhaust or re-cooled. The reject temperature ranges between 80° and 85°F wet-bulb, depending on governmental regulation, air velocity, and the expected metabolic heat generation rate of workers. Specifying the reject temperature is one of the first steps in planning air-conditioning systems. The ventilation engineer must be prepared to justify the reject temperature to management because of the economics involved. If too high, worker productivity, health, safety and morale will suffer. If too low, capital and operating costs become excessive.
• Critical Ventilation Depth. This mine-specific term is the depth at which the air temperature in the intake shaft has risen to the reject temperature through auto-compression and shaft heat loads. Work areas below the critical ventilation depth rely 100% on air conditioning to remove heat. The critical ventilation depth is reached at about 8000 to 10,000 feet below the surface, depending on surface climate in the summer, the geothermal gradient, and shaft heat loads such as pump systems.
• Base Heat Load. Base heat load is calculated at an infinite airflow at the reject temperature passing through the work area. The temperature of an infinite airflow will not increase as air picks up heat.
• Actual Heat Load. Actual heat load is measured or calculated at the average stope temperature. It is always greater than the base heat load because the average stope temperature is lower than the reject temperature. More heat is drawn from the wallrock.
• Marginal Heat Load. This is the difference between base and actual. It is the penalty paid for using less than an infinite airflow (i.e., the lower the airflow, the lower the inlet temperature required to maintain the reject, and therefore the higher the heat load.)
• Temperature Dependent Heat Sources (TDH). These heat sources depend on the temperature difference between the source and air. Examples include wallrock, broken rock, and fissure water (either in a ditch or pipe).
• Temperature Independent Heat Sources (TIH). These heat sources depend only on the energy input to a machine or device after the energy required to raise the potential energy of a substance, if any, is deducted. Examples include electric motors, lights, substation losses, and the calorific value of diesel fuel combusted.
• Passive Thermal Environmental Control. Passive control separates heat sources from ventilating airflows. Examples include insulating pipes and wallrock, and blocking off inactive areas.
• Active Thermal Environmental Control. Active control removes heat via airflow and air conditioning quick enough so that air temperature does not rise above the reject.
• Positional Efficiency. Positional efficiency is an important design parameter for mine cooling systems. It is the cooling effect reaching the work area divided by the machine evaporator duty. The further the distance between the machine and work area, the more heat that the cooling medium, air or water, picks up en route.
• Percent Utilization. Percent utilization is the ratio of the evaporator duty of the refrigeration plant over a year’s time in energy units to the duty if the plant had worked the entire year at 100% load. This consideration becomes important when evaluating surface vs. underground plants.
• Coefficient of Performance (COP). The usual definition is the evaporator duty divided by the work of compression in similar units. A modification for mine systems is the overall COP. It is the evaporator duty divided by all power consuming devices needed to deliver the cooling to the work sites. This includes pumps and fans as well as refrigeration machine compressors.

For those unfamiliar with mining terms, the following apply:

• Shaft. A shaft is a vertical opening or steep incline equipped with skips to hoist the ore, and cages (elevators) to move personnel and supplies. Electric cables and pipes for fresh water, compressed air, cooling water, and pump water are installed in shafts.
• Drift. A drift is a horizontal opening often referred to as a tunnel. However, a tunnel opens to daylight on both ends. Drifts, deep underground, do not.
• Stope. A stope in metal mining is a production site where ore is actually mined. In coal mining, coal is usually produced either by longwall or room-and-pillar methods.

SOURCES OF HEAT ENTERING MINE AIR

Projecting mine cooling requirements commences after mining methods, work sites, production rates, and equipment selections have been specified through the time frame for which the ventilation and cooling systems are expected to provide an acceptable work environment. This time frame is normally ten years, although it can vary depending on circumstances.
Major sources of heat or temperature increase include adiabatic compression of intake air flowing down shafts, electromechanical and diesel equipment, hot fissure water, wallrock and broken rock. Minor sources include pipes containing hot compressed air and hot water, oxidation of timber and sulfide minerals, blasting, and body metabolism. In the discussions to follow, heat flow is expressed in Btu/min. One ton of refrigeration removes 12,000 Btu/h of heat.

Adiabatic Compression

Adiabatic compression is not technically a heat source because heat does not flow from one substance to another. However, the process does result in a temperature increase for air flowing down a shaft, due to the conversion of potential energy into internal energy. Air descending a shaft also increases in pressure due to the mass of air above it.
For dry air at standard conditions (59°F at 14.696 psia), the specific heat at constant pressure, Cp, is 0.24 Btu/lb·°F. Cp can be assumed constant for most ventilation work, but the ventilation engineer should be aware that extreme conditions might warrant a closer calculation of Cp. When airflow descends, one Btu is added to each pound for every 778 feet. Or, one Btu of internal energy is removed for the same elevation increase on the exhaust side. The dry-bulb temperature change is 1/(0.24 x 778 x 1) = 0.00535°F per foot, or 1°F per 187 ft of elevation. The specific heat for water vapor is 0.45 Btu/lb·°F. So, for constant air-vapor mixtures, the change in dry-bulb temperature is (1 + W)/ (0.24 + 0.45W) per 778 feet of elevation, where W is the humidity ratio in pounds of water vapor per pound of dry air.
The wet-bulb temperature lapse rate is of most interest to mine ventilation engineers. The rate varies, depending on the entering temperature and humidity ratio, and the pressure drop in the shaft. It averages about 2.5°F wet-bulb per 1000 feet, and is much less sensitive to evaporation or condensation than the dry-bulb.
Theoretically, the heat load imposed on intake air due to adiabatic compression is given in Equation 1, which is a simplified form of the general energy equation.

1) Heat Load = (Airflow, ft³/min)(60 min/hr)(density, lb/ft³)( 1 Btu )(elevation change, ft)
778 ft-lb

Example: What is the equivalent heat load due to the adiabatic compression of 300,000 cfm @ 0.070 lb/ft³ density flowing down a 5000 ft deep shaft?

Solution: Heat Load = (300,000 ft³/min)(60 min/hr)(0.070 lb/ft³)( 1 Btu )(5000 ft)
778 ft-lb
= 8,097,686 Btu/h

The adiabatic compression process is seldom adiabatic however. Autocompression is a more fitting term. Other heating or cooling sources, such as shaft wallrock, the introduction of groundwater or water sprayed in the shaft to wet the guides, compressed air and water pipes, or electrical facilities, often mask the effects of adiabatic compression. The actual temperature increase for air descending a shaft does not usually match the theoretical adiabatic temperature increase, due to the following:

• The effect of seasonal and daily surface temperature fluctuations such as cool night air on the rock or shaft lining. Rock exhibits a thermal inertia effect, which absorbs and releases heat at different times of the day.
• The temperature gradient of the rock related to depth.
• Evaporation of moisture within the shaft, which suppresses the dry-bulb temperature rise while increasing the moisture content of the air.

Electromechanical Equipment

Electric motors and diesel engines transfer heat to the air. The loss components of substations, the electric input to devices such as lights, and all energy used on a horizontal plane appear as heat added to the mine air. Energy expended in pumps, conveyors and hoists to increase the potential energy of a material, after losses are deducted, does not appear as heat.
Vehicles with electric drives such as scoop-trams, trucks and electric-hydraulic drill jumbos release heat into the mine environment at a rate equivalent to the nameplate and a utilization factor. For example, a 150-hp electric loader being operated at 80% of nameplate for twelve hours per day would liberate (150 hp)(42.4 Btu/min·hp)(0.80)(12 hr/day)(60 min/hr) = 3,663,360 Btu/day. Dividing by 24 hours/day gives an average heat load over the day of 152,640 Btu/h. During the 12 hours the loader is operating, the heat load is doubled to 305,280 Btu/h. The dilemma for the ventilation engineer is that if heat loads are projected at the 152,640 rate, the stope temperature will exceed the reject for half of the day, and the stope will be overventilated for the other half. If projected at 305,380 Btu/h, the stope will be greatly overventilated when the loader is not present. Current practice is to put up with the additional heat load while the loader is present. The operator gets some relief when he leaves the heading to dump the rock. During this time, the ventilation system can partially purge the heading.
For diesel equipment, about 90% of the heat value of the fuel consumed, or 125,000 Btu/gal, is dissipated to the air as heat (Bossard, 1982). The heat flow rate is about three times higher for a diesel engine than for an equivalent electric motor. If the same 150-hp loader discussed above were diesel-powered instead of electric, the heat would average about 458,000 Btu/h over the day, and 916,000 Btu/h during actual loader operation. Both sensible and latent heat components of the air are increased since combustion produces water vapor. If a wet scrubber is used, exhaust gases are cooled by adiabatic saturation and the latent heat component increases even further.
Fans raise the temperature of the air about 0.45°F per inch of water static pressure (1999 ASHRAE Applications Handbook, page 26.1). Pressures up to 10″ water column are common in mine ventilation. This is detrimental only when fans, primary or auxiliary, are located on the intake side of the work areas or circuits.

Groundwater

Transport of heat by groundwater is the largest variable in mine heat loads, ranging from essentially zero to overwhelming values. Groundwater usually has the same temperature as the virgin rock. If hot drain water flows in an uncovered ditch, ventilating airflows can easily pick up more heat from the water than from wallrock. Thus, hot drain water should be stopped at its source, or contained in pipelines or in covered ditches. The pipelines could be insulated, but the main goal is isolating the hot water so that evaporation cannot occur.
Heat release from open ditches becomes more significant as airways get older and the flow of heat from the surrounding rock decreases. In one Montana mine, water in an open ditch was 40°F cooler than when it flowed out of the wallrock; the heat was transferred to the air. If this water were contained in a pipe, heat transfer to the air would be significantly less.

Example: Say that 20 gpm leaks from a rock fissure at 125°F. If the water enters the shaft sump at 85°F, what is the rate of heat being transferred to the air?

Solution: Heat rate = (20 gpm)(60 min/hr)(8.33 lb/gal)(1 Btu/lb·°F)(125° – 85°F)
= 399,840 Btu/h

Wallrock Heat Flow

Wallrock is the main heat source in most deep mines. Wallrock heat flow is unsteady-state; it decays with time due to the insulating effect of cooled rock near the rock-air boundary. Equations have been worked out for both cylindrical and planar-shaped openings. Cylindrical equations are presented here (Goch & Patterson, 1940). The method has two choices for the user: the instantaneous or the average heat flux rate. The instantaneous rate is recommended because it is better used for older tunnels or drifts. For young drifts, a series of instantaneous rates over short time periods is equivalent to average rates.
Temperature at the earth’s core has been estimated at about 10,300°F (Encyclopaedia Britannica, Fifteenth Edition, Vol. 17, page 595-601). Heat flows from the center of the earth to the surface at an average 0.07 W/m² rate. The implication for mine engineers is that a geothermal gradient exists. Rock gets progressively warmer as the mine deepens. The actual gradient varies from approximately 0.5°F to over 4°F per 100 ft of depth, depending on the thermal conductivity of the local rock type. Table 1 gives depths and maximum rock temperatures for a number of mining districts. Table 2 gives thermal conductivities and diffusivities for rock types commonly found in mining. These two variables are required for wallrock heat flow analysis.

Table 1 Maximum Virgin Rock Temperature
(Fenton 1972)

Mining District Depth, ft Temperature, °F
Kolar Gold Fields, India 11,000 152
South Africa 12,000 125 to 135
Morro Velho, Brazil 8,000 130
N. Broken Hill, Australia 3,530 112
Great Britain 4,000 114
Braloroe, BC, Canada 4,100 112.5
Kirkland Lake, Ontario 4,000 to 6,000 66 to 81
Falconbridge Mine, Ontario 4,000 to 6,000 70 to 84
Lockerby Mine, Ontario 3,000 to 4,000 67 to 96
Levac Borehold (Inco), Ontario 7,000 to 10,000 99 to 128
Garson Mine, Ontario 2,000 to 5,000 54 to 78
Lake Shore Mine, Ontario 6,000 73
Hollinger Mine, Ontario 4,000 58
Creighton Mine, Ontario 2,000 to 10,000 60 to 138
Superior, AZ 4,000 140
San Manual, AZ 4,500 118
Butte, MT 5,200 145 to 150
Homestake Mine, SD 8,000 134
Ambrosia Lake, NM 4,000 140
Brunswick #12, New
Brunswick, Canada 3,700 73
Belle Island Salt Mine, LA 1,400 88

Table 2 Thermal Properties of Rock Types
(by permission of Mine Ventilation Services Inc., Fresno, CA)

Thermal Conductivity Diffusivity
Rock Type (Btu/hr·ft·°F) (ft²/hr)
Coal 1.27 0.050
Gabro 1.37 0.092
Granite 1.11 0.129
Pyritic Shale 2.11 0.078
Quartzite 3.18 0.090
Sandstone 1.14 0.065
Shale 1.38 0.035
Rhyolite 2.00 0.043
Sudbury Ore 1.50 0.049
North Idaho Metamorphic 2.95 0.109

The Goch & Patterson method uses the following variables and equations for instantaneous heat flux:

Fo = Fourier number, dimensionless
 = Function of the Fourier number for the instantaneous rate, dimensionless
k = Thermal conductivity of rock, Btu/hr·ft·°F
 = Thermal diffusivity of rock (equals k/·C), ft²/hr, where
 = Rock density, lb/ft³
C = Heat capacity, Btu/lb·°F
θ = Average age of section, hours.
L = Length of section, ft
A = Cross-sectional area of section, ft²
P = Perimeter of section, ft
r = radius of circular section, ft, or equivalent radius of rectangular section; r = (A/)1/2
tvr = Virgin rock temperature, °F
ta = Air dry-bulb temperature, °F

These variables are used in the following equations for the instantaneous rate. The equations are easily PC-programmable, which is very important because a large number of sensitivity simulations are usually needed.

2) Fo = θ
r²

3)  = [1.017 + 0.7288log10(Fo) + 0.1459(log10(Fo))² – 0.01572(log10(Fo))³ –
0.004525(log10(Fo))4 + 0.001073(log10(Fo))5]-1 (Whillier & Thorpe, 1982)

4) Heat Flux (Btu/hr·ft²) = k(tvr – ta)()
r

5) Total Heat Flow (Btu/h) = (Heat Flux)(L)(P)

Example: A 500 ft long section of drift, 12 ft high x 15 ft wide, was driven in quartzite with a virgin rock temperature of 110°F. The drift was started 20 days before the face was reached, and the face is one day old. One design criterion is keeping the average dry-bulb temperature of the air in the drift at 80°F. How much heat will flow into the section?

Solution: From Table 2, the thermal conductivity of quartzite is 3.18 Btu/hr·ft·°F and the diffusivity is 0.090 ft²/hr. The average age of the section is (20 + 1 days)/2 = 10.5 days, or 252 hours. The cross sectional area of the drift is 12 x 15 = 180 ft² and the perimeter is (12 + 15) x 2 = 54 ft. The equivalent radius of the drift is (180/)1/2 = 7.57 ft. The following equations are then applied:

Using Equation 2, Fo = θ = (0.090)(252) = 0.396
r² 7.57²

Using Equation 3,  = 1.336

Using Equation 4, Heat Flux = k(tvr – ta)() = (3.18)(110 – 80)(1.336) = 16.84 Btu/hr·ft²
r 7.57

Using Equation 5, Total Heat Flow = (Heat Flux)(L)(P) = (16.84)(500)(54) = 454,700 Btu/h

Therefore, in order to keep the average temperature of the drift section at 80°F dry-bulb, 37.9 tons of refrigeration are required to remove the 454,700 Btu/h of heat.

Comments on the Use of the Goch & Patterson Method

The Goch & Patterson method lacks a convective heat transfer coefficient at the rock-air boundary. The method will overestimate heat transfer in a dry drift by a reported 8-15%. Nor does the method have a wetness factor. Since a drift with water on the perimeter will draw more heat from the wallrock, the method will underestimate heat flow. Almost all drifts have some degree of wetness on the floor, back and sidewalls, whether it is evident to the eye or not. Comparisons of the Goch & Patterson method with actual measurements and commercial software under typical conditions (a drift with 20% to 60% of the perimeter wetted) indicate that the overestimate due to lack of a convective heat transfer coefficient is close to a trade-off with the underestimate due to the lack of a wetness factor. When using Goch & Patterson for drift heat loads, the ventilation engineer should keep the drift section lengths under 200 feet, and should not apply any contingency factor to the calculated heat load.
For stoping, heat load calculations can be exasperating due to the number of variables involved. Irregular shapes, sporadic advance rates, intermittent TIH sources, fissure water, and nonhomogeneous or anisotropic rock are hard to model (Marks & Shaffner, 1993), (Duckworth & Mousset-Jones, 1993). For cut-and-fill stoping with a sand floor, measured heat loads are about 70% of the heat loads predicted by Goch & Patterson. Other stoping methods such as room-and-pillar or tabular reef mining are more amenable to planar heat load equations. Empirical graphs relating heat load to productivity and depth have been published (Patterson, 1992).
Ventilation engineers assigned the task of projecting heat loads for new mines or extensive tunnel projects can write their own personal computer program using the Goch & Patterson equations. Or, they could consider using one of the available commercial software programs. These programs account for convective heat transfer, wetness, elevation changes, and TIH sources that, when taken in combination, can make hand calculations tedious. However, program input must be carefully derived or the output will be misleading.
Heat from Broken Rock

Freshly blasted broken rock can liberate significant amounts of heat in a confined area. An estimate must be made of the broken rock’s initial and final temperatures, and where the rock is cooled en route from the face to the hoisting facility.

6) Heat (Btu) = (mass)(specific heat)(virgin rock temperature – final temperature)

7) Heat load (Btu/h) = Heat
(time, hr)

Example: A 12 ft high by 15 ft wide by 10 ft long drift round is blasted in quartzite where the virgin rock temperature is 120°F. Quartzite has a 168 lb/ft³ density and a 0.2 Btu/lb·°F specific heat. By the time the rock is hoisted to the surface four hours later, is has cooled off to 90°F. What is the heat load imposed on the drift and shaft?

Solution: Heat = (12 ft x 15 ft x 10 ft x 168 lb/ft³)(0.2 Btu/lb·°F)(120° – 90°F)
= 1.8144 x 106 Btu

Heat load = 1.8144 x 106 Btu = 453,600 Btu/h
(4 hr)

Heat from Minor Sources

The heat produced by oxidation of timber and sulfide minerals can be locally significant. This oxidation can, and has, caused mine fires. Fortunately, timber is seldom used for ground support in modern mines. The heat produced by blasting can also be appreciable. The typical heat potential in various explosives is similar to that of 60% dynamite, about 1800 Btu/lb (1999 ASHRAE Applications Handbook, page 26.3). This heat is usually swept out of the mine between shifts and thus is not tallied in heat load projections. Body metabolism is more of a concern in refuge chambers and is rarely if ever included in heat load projections. Although minor heat sources are usually neglected, the ventilation engineer must remain vigilant for cases where local effects might be significant.

Summation of Mine Heat Loads

Total heat load for a mine or mine section is essentially the summation of all TIH and TDH sources. It helps to plot heat sources on a schematic. The heat load from the surface to the entrance of the stope is assessed first, starting with the TIH sources since they influence TDH sources. Shaft heat loads and autocompression, and drift heat loads are added to the air en route to stope sites. The process should take only one iteration, resulting in a stope entering temperature. Then the stope heat load is calculated by assuming that the wet-bulb leaving the stope equals the design reject temperature. The air temperature entering the stope is guessed, and heat load equations are used to calculate the exit temperature. If this exit temperature exceeds the reject, a lower stope entering temperature is assumed, and a new exit temperature is calculated. The process is repeated with new stope entering temperatures until the calculated stope exit temperature equals the design reject.
If the entering stope temperature calculated from the surface is greater than the entering stope temperature calculated from the reject, a higher airflow, or air conditioning, will be needed. Psychrometrics can determine the magnitude of the airflow increase or cooling required.

HEAT EXCHANGERS

Underground heat exchangers can be water-to-refrigerant, air-to-refrigerant, water-to-water, air-to-water, or air-to-air. A brine can be used instead of water where freezing might occur. Heat exchangers can be direct, as with spray chambers, or indirect, as with conductive heat transfer through tubes or plates.
When designing large cooling plants for mine duty, follow typical ASHRAE guidelines for surface industrial plants. See the Air Handling Equipment and General Components sections of the 2000 ASHRAE Handbook – Systems and Equipment.

Shell-and-Tube and Plate Heat Exchangers

Shell-and-tube heat exchangers are the mainstay of refrigeration machines deployed in mining. Machines in the 200-400 ton range may use either direct expansion (DX) or flooded evaporators. In both cases the working fluid, refrigerant or water, is circulated through the tubes.
South African mines have made a strong case for using plate-and-frame evaporators in large surface chilled water plants (van der Walt, 1988). These machines can cool water down to within a degree of freezing without danger of rupture. In contrast, shell-and-tube evaporators should not be expected to chill water below 38°F. Manufacturers must be consulted.
Shell-and-tube water-to-water heat exchangers have been used in mine cooling systems to avoid pumping return water against high heads (the U-tube effect). Chilled water from the surface is sent down to the high-pressure (tube) side of the exchanger. Water on the low-pressure side operates district chiller systems or spot-coolers. The shell-and-tube method has never really caught on, perhaps due to the high-pressure supply and return piping. The Second Law limits the approach of the outlet high-pressure water temperature to the inlet low-pressure water temperature. This tends to limit heat removal in deep mines that would require at least three heat exchanger stations, in series, in the shaft.

Cooling Coils

Cooling coils can be direct-expansion (DX) or chilled water coils. DX coils are used with spot-coolers and typically range from 15 to 60 tons. Some modern spot-coolers use dual coils in parallel for compactness. Chilled water coils for use in district chiller systems also come in a wide range of sizes.
Air side fouling is the main operational problem with cooling coils in mines. Caution should be exercised in purchasing coils with fin spacing tighter than 6 fpi. Water side fouling is minimal if the water is of fair quality, the circuit is closed, and a corrosion inhibitor is added to the circuit.

Small Spray Chambers

Small spray chambers can be used as an alternative to cooling coils. Heat transfer is direct, air-to-water. Maintenance of a spray chamber is minimal. The amount of water sprayed is typically one-half to one-third that required for a cooling coil at the same duty. Some washing effect also occurs in the chamber.
That said, spray chambers are open systems where the water is dumped into the ditch or a collection pond after being sprayed. This water drains to the dewatering system or is pumped back to the chiller plant. Small spray chambers are still popular for small duties in mines that chill service water. But, the pumping system must be able to handle the increased service water requirement.
Cooling Towers

When heat loads are large, the full potential of a mine’s heat removal system will probably be needed. A key component of that heat removal system is the exhaust air. The ventilation engineer at a deep, hot mine must become proficient at designing underground cooling towers for condenser heat rejection, and spray chambers for cooling airflows.
Mining engineers do not use the standard HVAC&R method for assessing cooling tower performance, as described in the 2000 ASHRAE HVAC Systems and Equipment Handbook, Chapter 36. The following describes the South African factor of merit method developed in the 1970s for designing direct-contact heat exchangers (Whillier, 1977, Bluhm, 1981, and Burrows, 1982). The variables and equations are listed below. In addition, a psychrometric program is needed (see the 2001 ASHRAE Fundamentals Handbook, Chapter 6).

M = mass flow rate, water or air, lb/min.
q = heat rate to be transferred in the chamber, Btu/min.
h = enthalpy of moist air, Btu/lb.
S = Sigma energy of the air, Btu/lb; the total enthalpy minus the enthalpy of the liquid water
evaporated into the air (approximated by (Cpw)(W)(t), where t is the wet-bulb). Sigma
energy is dependent only on the wet-bulb and barometric pressure.
Cpw = Specific heat of water at constant pressure, = 1 Btu/lb·°F.
W = humidity ratio of moist air, lb water per lb air.
t = temperature, °F.
w = subscript denoting water.
a = subscript denoting air.
i = subscript denoting inlet.
o = subscript denoting outlet.
nw = water efficiency, dimensionless.
R = Tower capacity factor, dimensionless. This important number is the ratio of the heat
capacity of the water to the heat capacity of the air under the limits imposed by the
Second Law.
N = Number of transfer units, an intermediate factor for calculating water efficiency.
F = Factor of Merit, roughly equivalent to the UA factor in conductive heat transfer,
dimensionless. F ranges from 0 to 1. At 0, no heat transfer takes place. At 1, as much
heat transfer as permitted by the Second Law takes place.

8) S = h – (Cpw)(W)(t) Note that (t) is either the wet-bulb temperature for air, or the water
water temperature, depending on whether Sai or Swi is calculated.

9) q = (Mw)(Cpw)(twi – two)

10) q = (Ma)(Sao –Sai)

11) nw = twi – two
twi – tai

12) R = (Mw)(Cpw)(twi – tai)
(Ma)(Swi – Sai)

13) N = F
(1 – F)(R0.4)

14) nw = (1 – e-N(1 – R)) (for counterflow towers)
(1 – Re-N(1 – R))

Table 3 Factors of Merit
Mine Ventilation Practitioner’s Data Book
(Mine Ventilation Society of South Africa)

Factor-of-Merit Range
Vertical counterflow, open, unpacked 0.5 to 0.7
Single-stage horizontal crossflow 0.4 to 0.5
Two-stage horizontal crossflow 0.6 to 0.7
Three-stage horizontal crossflow 0.7 to 0.8
Commercial packed cooling counterflow tower 0.68 to 0.78
Commercial packed cooling crossflow tower 0.55 to 0.65

For designing an underground cooling tower, the designer must know the exhaust air mass flowrate, Ma, the wet-bulb and dry-bulb temperatures available at the tower, and the ambient barometric pressure. A psychrometric program must be available for calculating enthalpy, density, humidity ratio, and specific volume from the wet-bulb, dry-bulb and barometric pressure. Tower airflow is taken from measurements or from the mine plan. The temperature, if not measurable as for a new mine, can be assumed to approach the reject temperature. Experience has shown that air usually enters exhaust at about 82°-83° wet-bulb. The following sequence of design steps is followed:

Step
1. Calculate the heat rejection rate in the tower. This is the evaporator duty times the condenser heat rejection factor, typically between 1.2 and 1.4. Discuss with the manufacturer.
2. Select the condenser water flow and use Equation 9 to calculate tw in the tower, and in the machine condensers.
3. Specify the cooling tower diameter and the tower air velocity.
4. Calculate the Mw/Ma ratio.
5. Calculate the heat rejection rate per cfm in the tower.
6. Use Equation 8 to calculate Sai from the entering wet-bulb and barometric pressure.
7. Select a Factor-of-Merit for the tower using Table 3. Ensuring that this F can be met is discussed later.
8. Make a guess of the tower capacity factor, R. A good start is R = 0.5. Note, inputting
R = 1 into Equation 14 will result in division by zero. Proceed to Step 10 if R = 1, and use
the value of F for nw.
9. Use Equation 13 to calculate N, the number of transfer units.
10. Use Equation 14 to calculate the water efficiency, nw.
11. Use Equation 11 to calculate the inlet water temperature, twi.
12. Use Equation 8 to calculate Swi, the sigma energy of air at the inlet water temperature, twi.
13. Use Equation 12 to calculate a new tower capacity factor, R.
14. Compare the R calculated in Step 13 with the R guessed in Step 8. If different by more than 1%, return to Step 8 and re-guess R. Repeat Steps 9-14 until the R calculated in Step 13 falls within 1% of the R guessed in Step 8.
15. The air and water temperatures leaving the tower, and the evaporation rate, can now be calculated.

Steps 1 through 5 should be checked against the following experience-related design criteria before proceeding with Step 6:

• A realistic tw in the tower is 12° to 16°F.
• A realistic water loading in the tower is 6 to 18 gpm per ft².
• An optimum range of water velocity in machine condenser tubes, from 3 to 13 fpm, per manufacturer’s recommendations, is based on tubing material and water quality.
• A realistic maximum air velocity in the tower is 1,600 fpm.
• A range of realistic ratios of the mass flows of water to air, Mw/Ma, is 0.5 to 2.5.
• A realistic heat rejection rate in the tower is 32 to 65 Btu/h per cfm.

Values outside these design parameters can be used and not infrequently are (especially when plant duty is increased at a future date). But, the penalty paid is a higher condensing temperature and lower COP. Once q, F, Ma, tai, and Mw are specified, only one twi, two, and tao will balance all equations.

Example: Design a cooling tower for a 1000 ton refrigeration plant planned for deep hot mine. The exhaust airflow available for heat rejection is 250,000 cfm @ 83°F saturated. Barometric pressure is 15.226 psia (31″Hg, or 1000 feet below sea level). What size cooling tower is needed, how much condenser cooling water is required, what are the inlet and outlet air and water temperatures, and how much make-up water is needed?

Solution:

Step 1: For a refrigeration plant to produce 1000 tons® of cooling, it must reject about 1000 tons x 12,000 Btu/h·ton x 1.25 condenser heat rejection factor = 15,000,000 Btu/h.

Step 2: Select a condenser water flow. For this example, start with 1 gpm per 12,000 Btu/h rejected. The condenser flow is therefore 1250 gpm, or 10,413 lb/min at 8.33 lb/gal. The change in water temperature is calculated from Equation 9:

tw = 15,000,000 Btu/h = 24.0°F
(10,413 lb/min)(60 min/hr)(1 Btu/lb·°F)

That exceeds the realistic 12 to 16°F tw, so arbitrarily increase the water flow to 2,000 gpm and re-calculate tw. In practice, selecting the condenser water flow is anything but arbitrary. Generally, the higher the flow, the better. But, higher flows require larger condensers to keep tube velocity within design limits, larger cooling towers with more nozzles, and significantly larger pumps. The Law of Diminishing Returns takes effect quickly. Actual condenser water flow is a compromise between machine and tower performance, capital cost, and overall plant COP (operating cost). At 2000 gpm for this example, tw = 15.0°F, which is acceptable.

Step 3. Specify the cooling tower diameter by using a mid-range value of 12 gpm/ft².

(/4)d² = 2000 gpm  d =14.57 ft, say 15 ft.
12 gpm/ft²

Air velocity = 250,000 cfm = 1,415 fpm (< 1,600 fpm, so it is acceptable)
(/4)(15²)

Step 4: Calculate the Mw/Ma ratio.

Mw = (2000 gpm)(8.33 lb/gal) = 16,660 lbw/min

The specific volume for 83° saturated inlet air @ 31″Hg is 13.71 ft³/lba. Ma is therefore:

Ma = 250,000 cfm = 18,235 lba/min
13.71 ft³/lba

Mw/Ma = 16,660 = 0.914 ( 0.5 < Mw/Ma < 2.5, so it is acceptable)
18,235

Step 5: The heat rejection rate in the tower = 15,000,000 Btu/h = 60 Btu/h per cfm
250,000 cfm

This rate is approaching the upper acceptable limit. Consideration should be given to routing more air through the tower if possible. All of the design criteria have now been met.

Step 6: Sai @ 83° wet-bulb and 31″ Hg = 45.95 – (1)(0.0237)(83) = 43.98 Btu/lb.

Step 7: Select a factor-of-merit for the tower. From Table 3, an open, unpacked, vertical counterflow cooling tower can conservatively be expected to have a 0.55 factor-of-merit. If the tower is well designed and actually has a higher factor, the tower will return cooler water to the plant and COP will increase.

Step 8: Guess R = 0.5 (first pass).

Step 9: Calculate N from Equation 13. N = (0.55) = 1.613
(1 – 0.55)(0.50.4)

Step 10: Calculate nw from Equation 14. nw = (1 – e-1.613(1 – 0.5)) = 0.713
(1 – 0.5e-1.613(1 – 0.5))

Step 11: Calculate twi from Equation 11 (after manipulation, and assuming that twi – two = Δtw):.

twi = Δtw + tai = 15° + 83° = 104.04°F
nw 0.713

Step 12: Calculate Swi @ 104.04° and 31″ Hg = 77.11 – (1)(0.047)(104.04) = 72.21 Btu/lb.

Step 13: Calculate the new R using Equation 12.

R = (16,660)(1)(104.04 – 83) = 0.681
(18,235)(72.21 – 43.98)

Step 14: The new R, 0.681, is higher than the 0.5 R guessed in Step 8. Return to Step 8 and make a second guess. Keep iterating until the R calculated in Step 13 equals the R guessed earlier in Step 8. This occurs at R = 0.662.

Step 15: All other values can now be calculated.

Per Step 11, twi = 106.09°F.

two = 106.09° – 15° = 91.09°F.

Sao = Sai + 15,000,000 Btu/h = 43.98 + 13.71 = 57.69 Btu/lb
(18,235 lb/min)(60 min/hr)

tao = 94.5°F (via psychrometric iteration)

The water evaporated in the tower is the difference in humidity ratios, W, times the mass flow of dry air. From psychrometric equations, W83° = 0.0237 lbw/lba and W94.5° = 0.0347 lbw/lba

Evaporation = (18,235 lba/min)(0.0347 – 0.0237 lbw/lba) = 24.1 gpm
8.33 lbw/gal

Total make-up water depends on the evaporation rate, the water carry-over if any, and the blowdown used to control dissolved solids in the condenser circuit. Leakages occur in some systems. If so, leakage (and carry-over) can be deducted from the blowdown. Make-up water is usually planned at 1-3% of the condenser water flow. It depends on the quality of the make-up water, the allowable cycles of concentration of dissolved solids, and the water treatment plan.
Vertical unpacked cooling towers in mines often use clog-resistant full-cone nozzles circling the top of the tower, at least 40 feet off the pond. South African mines prefer ham-type sprayers. Nozzle pressure of 30 psig is typically specified. Lower water pressures don’t generate the finer water droplets preferred for heat transfer while higher pressures increase pumping costs. Higher pressures can also impinge water drops into sidewalls where the water forms sheets that run down the sides. This drastically reduces the surface area of the water flow, which reduces heat transfer. Rings circling the tower are recommended to kick the water running down the sides back into the airstream. Unpacked towers don’t have as high a factor of merit as towers with film packing or splash bars, but they are virtually maintenance-free and offer low resistance to airflow. Figure 1 shows a typical underground vertical counterflow cooling tower.
After a cooling tower has been put into operation, the actual factor of merit should be determined. This is accomplished by measuring air and water flowrates, and temperatures, at the tower inlet and outlet. The cooling tower equations are then worked in reverse. After solving for the actual factor of merit, it can be used to determine performance at other inlet conditions. This applies for industrial and commercial cooling towers as well.

Large Spray Chambers (Bulk Air Coolers)

The procedure for designing spray coolers is the same as for cooling towers, with the following minor changes:

15) q = (Mw)(Cpw)(two – twi)

16) q = (Ma)(Sai –Sao)

17) nw = (1 – e-R(1 – X)) where X = eN for crossflow chambers.
R

A perfect counterflow tower has a factor of merit of 1.0. But, the factor of merit for a single crossflow chamber cannot exceed 0.63 (Bluhm, 1981). Two-stage crossflow chambers are most often specified. Counterflow performance is approximated and the counterflow equation for water efficiency can be used. Three-stage chambers can be designed when water flow must be limited to control pumping costs. Four-stage chambers are rarely, if ever, worth the extra cost.
Spray chambers often use Vee-jet nozzles at 30 psig. Nozzles are placed uniformly along the chamber length and are designed to cover the cross-section evenly. Sprays should just make it to the back of the chamber. Mist eliminators are usually installed at the chamber exit. Whereas cooling towers need make-up water to replace evaporated water, bulk air coolers make water via condensation. This water can be sent to the condenser side as make-up. Figure 2 shows a typical 2-stage horizontal crossflow bulk air cooler.

MINE COOLING TECHNIQUES

A mine cooling system is essentially a heat removal system. Basically, a substance such as air, water, or ice, is sent into the mine at a low enthalpy state and removed at a higher one. In deep, hot mines, heat is typically rejected to water being pumped to the surface, and to exhaust air being drawn from the mine. There are many combinations and variations on how this is accomplished. Economics and site-specific conditions determine the optimum methods.

Increasing Airflows

This alternative should be considered first when a mine starts to have heat problems. It is usually less expensive to moderately increase airflows than to install refrigeration if the mine is above the critical ventilation depth. Increasing airflows also helps remove diesel fumes, an increasingly important consideration for modern mining. However, in deep, usually older mines with small cross-sectional airways, airflow increases may not be practical due to the cube relationship between fan power and airflow increase through a given resistance. Circuit resistance reduction via new airways or stripping existing airways is very expensive.

Chilling Service Water

When a mine requires additional heat removal and airflow increases are not practical, chilling the service water should be considered. Most mining methods require that water be sprayed on rock immediately after blasting to control dust. Chilled water can intercept rock heat before the heat escapes into the air. This is a very flexible method of heat removal because it is applied when and where it is needed the most. After blasted rock is removed, the water is turned off and routed elsewhere. Main water lines should be insulated when service water is chilled.
Water is usually chilled in surface plants. Water is used in a single-pass system if regional supplies are plentiful. If scarce, it is pumped to the surface, recooled and returned underground. Recycling can ease discharge permit requirements and thus save on treatment costs. Some mines have zero discharge permits. Regions with low winter temperatures and low relative humidities during summer have a natural cooling capacity that is adaptable to water chilling on the surface. Warm mine water is pre-cooled in an evaporative cooling tower before being sent to the refrigeration plant.
Spot-cooling, especially with spray chambers, works much better if the coolers receive chilled water instead of warmer water.

Reducing Water Pressure and Energy Recovery Systems

All mines send water underground for drilling, cleaning, suppressing dust, and wetting down broken rock. Hot mines employing extensive cooling systems often send large volumes of water underground solely for air conditioning. The pressure of descending water must be periodically broken. The most common methods are the open cascade system and pressure-reducing valves. Turbines can also be used to break the pressure and to recover a significant portion of the potential energy that would otherwise be lost. Two types of turbines are suitable for mine use – the Pelton wheel and a centrifugal pump specially designed to run in reverse. The Pelton wheel is most often used. The rotor is shaped like the spokes of a wheel, with cups attached to the ends of the spokes. One or two nozzles shoot high pressure water on to the cups, thus spinning the wheel. It is at least 80% efficient over a range of flows, is simply constructed, and is readily controlled. A wide operating range is important because water demand fluctuates. A turbine can turn either a generator or pump. Generators are most often used. This separates the service and cooling water from the mine dewatering system so that downtime in one system has less chance of disrupting the other.
Besides providing power to help return service water to the surface, turbines have another advantage. Unrecovered potential energy is converted to heat at a rate of 1 Btu/lb per 778 ft of depth. If, for example, a 6000-foot deep mine uses 1000 gpm for air conditioning without energy recovery, the water will heat up by 7.71°F. If 80% efficient turbines are used, the water temperature rise is about 0.2 x 7.71° = 1.54°F. The refrigerating effect lost is only 64 instead of 321 tons.
Other energy recovery devices besides turbines are hydro-transformers (a large piston device that transfers the force from the high pressure side to the lower pressure side), and three-pipe feeder systems that can actually deliver chilled water on one side while pumping out crushed ore on the other. These concepts have undergone testing in Europe and South Africa.

Bulk Cooling vs. Spot-Cooling

Ventilation engineers must decide on the proper balance between bulk cooling and spot cooling. Bulk cooling via a centrally-located plant cools the entire mine, or a large section of the mine. Benefits are lower cost per ton installed, generally better maintenance, and lower temperatures in non-stoping areas such as haul drifts. Bulk cooling intake air is often done in warm-climate mines to provide winter-like or better conditions year round. Air is cooled in large direct-contact spray chambers adjacent to the shaft and then is injected into the shaft below the main landing.
Cooling the entire mine draws more heat from surrounding wallrock, so a larger system must be designed in order to ensure proper stope cooling. In other words, positional efficiency suffers. When a multi-level mine is bulk-cooled, cooling may be wasted on upper levels where heat load is low.
Spot-cooling, on the other hand, provides adequate temperature control in exploration and development headings, and in stopes on the fringes of mining activity. Total heat load is lower, but cost per ton is higher and temperatures in some areas might exceed design limits.

Combination, or Integrated Surface Systems

Combination systems can cool both air and water. Surface plants devote a higher fraction of cooling capacity to bulk cool intake air in the summer. In winter, a higher fraction is used to chill service or air-conditioning water. Water delivery underground can be accomplished via open or closed systems, with or without energy recovery. Figure 3 shows components of an integrated mine cooling system.

Underground Refrigeration

Besides spot coolers, larger refrigeration machines can be located underground. They usually produce chilled water for cooling air in spray chambers. Heat is rejected to exhaust air via cooling towers. Another method is to operate district cooling systems. A chiller produces water for a closed network of cooling coils installed in parallel. These coils can be used in auxiliary systems at individual work areas, or installed in a bank. As with spot-coolers, coils should be installed upwind of blasting to limit air-side fouling. Condenser heat from district chiller systems is rejected either to service water or to the mine dewatering system.

Ice Plants

For ultra-deep mines (>12,000 feet), or for mines already using existing water and airflows to the maximum for heat rejection, ice cooling should be considered. In going from 32°F to 90°F before being pumped out of the mine, cooling water starting as ice can remove about 4.5 times the heat as the same mass flow of chilled water in going from 45°F to 90°F, per the calculations below:

Heat removal (Btu/lb) = Sensible + Latent

Heat removal of chilled water = (1 lb)(1 Btu/lb·°F)(90° – 45°) = 45 Btu/lb

Heat removal of ice = (1 lb)((1 Btu/lb·°F)(90° – 32°) + 144 Btu/lb) = 202 Btu/lb

Heat removal factor increase of ice over water = 202/45 = 4.5 times

South African mines have been at the forefront in this application (Sheer, 2001). Basically, two methods are available; chunk and slurry. Both methods send ice to underground chambers where it is mixed with a portion of the warm water returning from the mining area. The cold mixed water is then sent back to the mining area.
Several successful systems have been installed to date. Cost has dropped as the technology improves and manufacturers recoup their R&D costs. The overall COP of ice systems for ultra-deep mines is now competitive with more traditional cooling methods.

Thermal Storage

This innovative technique developed in Canada uses near-surface ice stopes or rock rubble to effectively, and inexpensively, heat intake air in the winter and cool it in the summer (Stachulak, 1989).

Controlled Recirculation

This technique, used in conjunction with bulk air cooling underground, can reduce ventilation and air conditioning requirements in older, deep mines, especially when heavily mechanized (Tien, 1999). Besides increasing air velocities in work areas without drawing more surface air through a high resistance circuit, controlled recirculation reduces the heat load caused by autocompression. Using Equation 1, for every 100,000 cfm at standard density brought from the surface, the lost cooling capacity per 1000 feet of descent is:

Heat Load = (100,000 ft³/min)(60 min/hr)(0.075 lb/ft³)( 1 Btu )(1000 ft)
778 ft-lb

= 578,406 Btu/h, or 48.2 tons of cooling lost

 

Cabs and Vests

As noted earlier, mechanization can add significant heat to the mine environment, especially in confined auxiliary-ventilated spaces. Most non-coal mines have converted to diesel equipment during the last twenty years, although some employ electric loaders and trucks. Coal mines are also mechanizing, but more slowly. The biggest problem in coping with diesel engine heat is related to the greatest advantage of these vehicles: mobility. The heat and emissions from a diesel vehicle working in a confined area can tax almost any ventilation system. More and more, air-conditioned cabs are being specified for large diesel vehicles. Cabs come equipped with window-type air conditioners and HEPA filters to capture diesel particulate and dust. Cabs also protect the operator from excessive noise. After the diesel vehicle has left the heading, the mine ventilation and cooling system can provide an acceptable environment for drillers, blasters, bolters and other personnel. It should be noted that federal regulations concerning exposure to diesel smoke and noise are becoming stricter.
Cabs are expensive, adding about 12% to the purchase price of a 3.5-yard scoop-tram. However, a loader often visits multiple headings in the course of a week. The cost of a cab is less than maintaining the design reject wet-bulb temperature in all headings at all times.
Cooling vests are not popular in the mining industry. They are bulky, reduce mobility, and are time-consuming to prepare and use. South African mines have used dry-ice vests to acclimatize workers; a three to five day process. Vests have limited application for mechanics, electricians, pipe-fitters and others who must enter hot areas in order to set up the ventilation and cooling. Vests employing blue-ice packs last two to three hours. Vests using compressed air venturi-type coolers require an umbilical cord.

Other

Other methods in the conceptual stage include the air cycle (air compressed on the surface and sent underground to a turbine, turning a generator and exiting at -40°F), and the ammonia cycle (sending down liquid ammonia, evaporating it, and sending the vapor back to surface condensers). These methods may be best suited to ultra-deep mines where other cooling methods are already fully deployed.
Transferring heat from current stopes to the wallrock or rock rubble in previously worked out stopes has also been considered. This is the only method that does not remove heat from the mine. The refrigeration equipment would have to operate at a high condensing temperature to produce hot enough water to transfer heat to worked out stopes.

SELECTING A MINE COOLING METHOD

After mine cooling and ventilation requirements have been projected, the designer must analyze and select the best method(s) for meeting those requirements. Cost-benefit is the most widely used comparative technique, but hardware reliability, dependency on outside factors, flexibility, safety, and technological level, are just as important. Some factors that influence alternatives selection are:

• Seasonal Ambient Conditions. Warm climate mines tend to bulk cool air on the surface in industrial direct-contact spray chambers located close to the intake shaft.
• The Orebody and Mining Methods. The more massive the orebody, the more attractive bulk cooling becomes. When stopes are scattered and continuously advanced into new areas, district or spot cooling might be better.
• The Mining Rate. This is a critical point. Heat removal is energy-related, not necessarily power related. A fast mining rate prompts a high instantaneous heat load expressed in Btu/h, but less heat energy per ton of production expressed in Btu. This is because wallrock is covered by fill or isolated before the total heat energy of the wallrock has escaped into the airstream. It is the Btu per ton of production that incurs air conditioning costs. Leave as much heat in the wallrock as possible.
• Size and Condition of Major Airways. In older mines, small airways often limit the potential for airflow increases. This may prompt the need for air conditioning sooner than it normally would have been necessary.
• Heat Sources. The contribution of TIH and TDH sources to the total can help determine the balance of passive to active thermal environmental controls, the ratio of airflow to air conditioning, and whether or not cabs should be specified.
• Cost of Power, Water, Labor and Supplies. Knowing these costs is critical for assessing the economic optimum of capital expenditure to control operating costs. For example, if power cost is high, spending extra capital for a higher COP system may be warranted.
• Governmental Regulations. These can influence the size of the system due to heat stress standards. Certain safety issues are also impacted, such as not using combustible pipe insulation or ammonia machines underground.

Basic cooling alternatives for specific cases are summarized in Table 4. Airflows are described as limited, medium, or large. One way to express airflow for a given mine is the ratio, tons of airflow per ton of ore. The time variable drops out, and, being dimensionless, the ratio facilitates international comparisons. A limited airflow is defined as less than 8 tons of air per ton of production. A medium airflow is 8 to 16 tons per ton, and a large airflow is over 16 tons per ton. These ranges are based on an unpublished study conducted by the author for approximately 100 mines of all types, world-wide. The ranges discussed are for heat removal only. Additional airflow needed for methane, radon or diesel emissions removal must be addressed separately.

Table 4 – Basic Cooling Alternatives

Warm Climate Cool/Cold Climate
Massive Orebody Large or medium airflow Medium airflow
(Deep) Chill service water Chill service water
Bulk cool air on surface Bulk cool air underground
Bulk cool air underground Thermal storage

Massive Orebody Large airflow Large or medium airflow
(Shallow) Bulk cool air on surface Chill service water on surface
Chill service water Shell-and-tube
Shell-and-tube

Scattered Orebody Large airflow if not too deep Large or medium airflow
(Multi-Level) Chill service water Chill service water
Bulk cool air on surface District chiller systems
District chiller systems Thermal storage
Spot-coolers/spray chambers

Ultra-Deep Orebody Limited airflow Limited airflow
(Massive and/or Chill service water Chill service water
Multi-Level) District chiller systems District chiller systems
Ice cooling Ice cooling
Controlled recirculation Controlled recirculation
Possible exotic system Possible exotic system

Small Orebody Bulk cool air on surface District chiller systems
Chill service water Chill service water
District chiller systems Spot-coolers
Spot-coolers

Porous Rock District chiller systems District chiller systems
Spot-coolers Spot-coolers

MECHANICAL REFRIGERATION PLANTS

Surface Plants

Centrifugal or helical rotary screw machines are typically used in surface plants to chill water or bulk cool air. Details can be found in the 2000 ASHRAE HVAC Systems and Equipment Handbook. Banks of machines are usually installed in parallel. Plant design must accommodate one machine being down at any given time for maintenance while the others operate. Shell-and-tube heat exchangers are standard, although plate-and-frame are used if chilled water close to the freezing point is specified. The most common refrigerant to date, at least for positive-displacement compression, is HCFC-22. Ammonia is also commonly used in surface plants. Absorption machines can be considered if external waste heat is available.

Underground Plants

Large underground plants do the same work as surface plants, but they are closer to work areas. Better positional efficiency and percent utilization are the advantages. Whereas surface plants use atmospheric air for heat rejection, underground plants use mine exhaust air. Rejecting heat to exhaust air also raises the natural ventilation pressure, which assists circuit fans. Components for underground machines must be disassembled for transport down the shaft.
The main disadvantage of underground refrigeration is that heat rejection is limited by the amount of available exhaust air. Excavating underground refrigeration rooms and spray chambers is more costly than erecting pre-fabricated surface buildings. Maintenance is also more difficult due to shaft logistics. Power is more difficult to supply to an underground facility, and subject to more disruptions.

Spot Coolers

Spot coolers with 15 to 100 ton capacity permit the driving of long development headings, or the cooling of exploration sites prior to the installation of primary ventilation and cooling equipment. Development headings can be advanced more rapidly and under more comfortable conditions. Condenser heat is most often removed by service water although some air-to-air condensers are used.
Spot coolers use reciprocating or scroll compressors. Hermetic scroll compressors are becoming more popular because they can handle liquid slugging better than reciprocating compressors. They are also less expensive. Spot-coolers use DX air cooling coils. The packaged unit includes a fan, which draws air through the coil (or coils, as in a dual-coil unit) and then blows it through duct to the heading. Spot-coolers must be compact and portable since they are moved often. Electricians hook up the power, and fitters attach the service water pipe. The service water required is typically 1.1 gpm per ton. It can be less if the water temperature is under 55°F, or more if it is over 70°. A return drain pipe is recommended to prevent contact between the hot discharge water (often over 100°F) and the ambient air. Coils sometimes receive dusty air immediately after blasting (when miners are gone from the work area). If so, they must be washed at least every other day.
Spot-coolers are expensive, but often are the only choice for cooling exploration, development, and small-scale stoping on the fringes of mining activity.

Maintenance

Mines with extensive systems (say, a large chiller plant, or over 10 spot-coolers), should employ a mechanic specializing in refrigeration. Mines with over 2000 tons probably need a second mechanic. These persons should be factory-trained and must be certified to handle refrigerants. Refrigeration specialists can be assisted periodically or full-time by apprentice mechanics. Another viable approach is to enter into a maintenance contract with the manufacturer or supplier of the equipment, or an independent HVAC&R shop. Some mines have a full-time person cleaning coils.
A fouling factor (units of ft²·h·°F/Btu) should be calculated from lab analysis of the condenser water, especially for district chillers using sump water. Planning a tube-cleaning regimen, either manual, acid circulation, or automatic with brushes and a flow reversal valve, is critical. Underground condensers can plug off within a couple weeks without cleaning, depending on fouling factor. Water treatment is needed to control scale, corrosion, and biological organisms in surface or underground plants with cooling towers.

MINE AIR HEATING

General Considerations

Cold climate mines typically heat intake air in the winter. In Canada, heating intake air can cost more than all other ventilation costs combined (Hall, 1989). Without heat, water in the shaft will freeze, which can disrupt hoisting operations and damage shaft support members, cables and pipes. Very cold air and icy floors are safety and health hazards. The necessity for wearing heavy gloves and other protective clothing can make routine tasks difficult. Intake air is typically heated to just above the freezing point. Autocompression and shaft heat loads further temper the air as it downcasts into the mine.
Steam coils operated by boilers combusting wood, coal, fuel oil, or natural gas often served as shaft heaters in the past. Electric resistance heaters have also been used, although they are expensive to operate. Waste heat from compressor stations has also been applied.
When exhaust and intake shafts are located close together, a circulating glycol or heat pump system can be used to transfer heat from exhaust air to intake. The psychrometric chart shows that for every degree of total heat (sensible plus latent) given up by warm saturated exhaust air, the same mass flow of cold intake air can be heated sensibly by 4°. Either coils or a cooling tower extracts the heat from exhaust air. Then, coils are used to transfer this heat to the intake air.
Controlled recirculation, up to 25% of total airflow, can also be applied to heat intake air (Hall, 1989). The system is temporarily shut down during blasting times.
Some cold-climate mines remove the primary production shaft from the ventilation circuit. A slight upcast flow of uncontaminated air maintains good conditions in the shaft for hoisting ore and moving personnel and supplies. The disadvantage of this method is transferring the ventilation duties of the production shaft to one or more expensive stand-alone intake airways.
Natural gas and/or propane heaters are typically used at modern mines. Natural gas is preferred because it is less expensive and it burns cleaner. Where natural gas is not available, propane must be trucked to the mine site. The same heater can be easily set up to burn either natural gas or propane. Thus, propane can be used for back-up in case the natural gas is cut off. Direct-fired heaters are usually preferred because the entire calorific value of the fuel enters the intake airstream. If indirect heaters are used, roughly 15-25% of the heating value will be lost up the flue pipe.

Practical Design Considerations

Two types of natural gas or propane heaters have been used to heat intake air. The simplest is a grid of burner bars installed in a housing at the intake shaft. The housing can have louvers to adjust the flow of intake air, and to mix air from the heaters with outside air. The second is a crop dryer type of burner. Temperature sensors installed downstream can modulate both heater types. This ensures that no more heat is applied than necessary to bring the temperature of the mixed intake air to 34°F.
Carbon-monoxide sensors should also be installed downstream of the heaters. Experience at two western mines shows that the CO content of intake air heated by direct-fired burners can reach 15-20 parts per million.

Equation 18, below, is used to calculate the total heat required, assuming that the air has a low humidity ratio (which is the case for very cold air), and that no water is evaporated in the heater. Heating values for different fuels are given in Table 5.

18) Heat (Btu/h) = (Airflow, cfm)(60 min/hr)(density, lb/ft³)(0.24 Btu/lb·F°)(Δt, °F)

Where Δt = 34°F minus the intake air temperature

Table 5
Heating Values for Different Fuels

Natural gas …………………. 1,000 Btu per cubic foot (Kennedy, 1996, p.79)
Propane …………………….. 90,000 Btu per gallon (Kennedy, 1996, p. 79)
Bituminous Coal …… 12,300 to 14,400 Btu per pound (Abbeon Cal, 2001, p. 213)
Fuel Oil …………………… 143,000 Btu per gallon (Abbeon Cal, 2001, p. 213)
Wood ……………….. .. 15-31 million Btu per chord (Abbeon Cal, 2001, p. 213)

Example: A mine is located where the atmospheric air temperature can drop to -20°F for two or more weeks per year. Occasionally the temperature drops to -30°F. An intake shaft handles 400,000 cfm, and the density of the air entering the shaft in winter is 0.070 lb/ft³. It is desired to keep this shaft free of ice. What heating should be installed at the shaft inlet?

Solution: Sizing heaters is usually based on average cold periods, not extreme cold snaps. In this case, a direct-fired heater will be sized to raise -20°F air to 34F°. When the temperature drops to -30° for short periods, intake airflow should be temporarily reduced. Using Equation 18:

Heat = (400,000)(60)(0.070)(0.24)(34°F – (-20°F)) = 21,800,000 Btu/h

If natural gas is used, the volume required would be:

21,800,000 Btu/h = 21,800 cubic feet per hour
1000 Btu/ft³

If propane is used, the gallons required are:

21,800,000 Btu/h = 242 gallons per hour
90,000 Btu/gal

MINE VENTILATION

Mine ventilation supplies oxygen to underground facilities, and removes dangerous or harmful contaminants such as methane, radon gas, strata gases, dust, blasting fumes and diesel emissions. Ventilation also removes heat and helps control humidity at deep, hot mines. Planning a ventilation system consists of five basic steps:

1. Determining airflows.
2. Planning the primary circuit.
3. Specifying circuit fans and their installation.
4. Determining auxiliary system requirements.
5. Assessing health and safety aspects.

Determining the Airflows

Mining operations generate differing types and amounts of contaminants, and airflows dilute and remove these contaminants. The ventilation engineer must work closely with the mine planning staff to understand where and how much production will take place, and what contaminants will be generated. The federal Mine Safety and Health Administration (MSHA) regulates contaminant concentrations to limits specified in the Federal Register, CFR 30. Controlling the single-most problematic contaminant normally keeps all others within their legal limits. For coalmines, the contaminants of concern are typically methane and coal dust. For uranium mines, it is radon gas. For non-dieselized hardrock mines, it is usually silica dust and blasting fumes. For dieselized mines, it is typically diesel emissions. Design airflows for dieselized non-uranium metal mines range from 75 to 150 cfm per diesel horsepower, depending on the reference cited. With the present emphasis on controlling diesel particulate, the ventilation engineer should start planning at 100 cfm/hp.
Total airflow is a summation of airflows for individual work areas, plus a leakage factor. Leakage is defined as airflow that does not ventilate any active work area or permanent site such as a pumproom. A “tight” system minimizes leakage via well-constructed doors and seals, by minimizing the number of possible leakage paths, and by careful fan placement. Leakage can range from 10% of total airflow at a tight metal mine to 80% at some coalmines.
The ratio, tons of air per ton of production, described for Table 4, is about 2 to 4 for block cave mines, 6 to 8 for non-dieselized cut-and-fill metal mines, and 9 to 16 for dieselized metal mines. Gassy coalmines and uranium mines can have significantly higher ratios, depending on the methane or radon generation rate.

Example: A new mechanized cut-and-fill gold mine is planned. Ore production is expected to be 1,200,000 short tons per year. Intake air density is 0.070 lb/ft³. What is the rough airflow required for ventilating this mine?

Solution: The airflow range is 9 to 16 tons of air per ton of ore for dieselized metal mines. For a first pass guess, assume an average 12.5 tons per ton. The total weight of the air coursing through the mine in a year’s time would be:

(1,200,000 tons of ore per year) (12.5 tons air per ton ore) = 15,000,000 tons of air per year.

The airflow rate would therefore be:

Airflow (cfm) = (15,000,000 tons/year x 2000 lb/ton) = 815,400 cfm
(0.070 lb/ft³)(525,600 min/year)

Using such ratios provides a good first guess. However, the ventilation engineer should derive the total airflow by listing all operations and adding the specific airflows required to ventilate each operation (zero-base planning). As with specifying reject temperature for a hot mine, the total airflow selected should be justified to management since the economics involved are significant.
Airflow specification may change with time due to production, equipment or mining method changes.

Planning the Circuit

Once the airflow is specified and work sites plotted, the ventilation engineer must lay out the primary circuit. The three basic types of airways are: intake, those in the work areas, and exhaust. Sizing airways is normally based on keeping velocity within acceptable limits. If too low, the airway is oversized and thus costs more than necessary. If too high, pressure drop becomes large which raises operating costs. Velocities exceeding 1200 fpm in production shafts and haul drifts also create dust problems and can lead to employee discomfort. On the other hand, velocities in bare circular concrete exhaust shafts can approach 5000 fpm if necessary.

Resistance to airflow is calculated using Atkinson’s work, first published in 1854.

15) ∆H = RQ² d where ∆H = pressure drop, inches water gage (“WG)
0.075 R = resistance, “WG·min²/ft6
Q = airflow, cfm
d = actual air density, lb/ft³
0.075 = standard air density, lb/ft³

16) R = kLP where: k = friction factor, lb·min²/ft4 (k-factor includes the effects of
5.2A³ pipes and ground support as well as rock surface roughness)
L = length, ft
P = perimeter of opening, ft
5.2 = conversion factor, lb/ft²·”WG
A = area of opening, ft²

Example: Mine plans call for 65,000 cfm to be sent through 2000 feet of 10’W x 10’H drift. The k-factor from previous measurements of similar drifts is 50 x 10-10 lb·min²/ft4. The average temperature is 75° wet-bulb and 80° dry-bulb. The barometric pressure is 13.8 psia (28.10″Hg). What is the resistance of this drift, and what pressure drop will the air experience?

Solution: Using the psychrometric equations in the 2001 ASHRAE – Fundamentals Handbook, pages 6.12-6.14, the density is calculated at 0.0683 lb/ft³.

R = KLP = (50 x 10-10)(2000)(40) = 7.69 x 10-11 “WG·min²/ft6
5.2A³ (5.2)(100³)

H = RQ² d = (7.69 x 10-11)(65,000²) 0.0683 = 0.30″WG
0.075 0.075

A mine ventilation circuit contains airways in series and parallel combinations. The overall resistance of airways in series and in parallel (Hartman, 1997) is:

17) For series: RT = R1 + R2 + R3 ּּּּּּ + Rn

18) For parallel: 1 = 1 + 1 + 1 + ….. + 1
RT R1 R2 R3 Rn

Example: If Airway #1 has a resistance of 1 x 10-10, Airway #2 has a resistance of 2 x 10-10 and Airway #3 has a resistance of 3 x 10-10, what is the resistance of these three branches in series? in parallel?

Solution: Series: RT = 1 x 10-10 + 2 x 10-10 + 3 x 10-10 = 6.0 x 10-10 “WG·min²/ft6

Parallel: 1 = 1 + 1 + 1
RT (1 x 10-10) (2 x 10-10) (3 x 10-10)

RT = 1.92 x 10-11 “WG·min²/ft6

Modern ventilation network analysis uses high-speed digital personal computers. The computer uses Kirchhoff’s laws, given below, to balance airflows:

1. The summation of airflows into a junction equals the summation out.
2. The summation of pressure drops around any enclosed mesh equals zero.

Most network programs use a balancing algorithm based on the work of Hardy Cross in the 1960s-1970s. The program iterates as it converges on final balanced airflows. Fan curves or regulators can be inserted in almost any branch. Computer simulation permits quick analysis of a wide range of scenarios. Although ventilation planning was accomplished surprisingly well in pre-computer days, no modern engineer would design ventilation systems without a network program. The earliest public domain mainframe programs available in the U.S. were written at Penn State University and the Michigan Technological University. These programs, and others, are now available for the PC.
Regulators or section booster fans control the airflow in branches. Without regulation, too little or too much airflow may occur. Despite that fact, circuits should be designed with as many free-split branches (branches without a fan or regulator) as possible. This helps minimize overall resistance. Free-split branches are often located in the circuit extremities.
Ideally, a mine should have more intakes than exhausts. Safety is enhanced, because the miners have more choices for escape, and because there are more paths to bring in fresh air, in case a fire occurs in one of the intakes. A fire can block the escape path to a particular shaft. Also, exhaust shafts can handle greater air velocities and hence larger quantities, so fewer exhaust shafts are needed.
Metal mines often contain circuit booster fans. Underground boosters can create neutral points in the system where air short-circuits from intake to exhaust above the point, and recirculates from exhaust to intake below the point. Uncontrolled recirculation should be minimized as much as possible.
Exhausting primary circuits are commonly used at both metal and coalmines. Intakes are thus unencumbered by airlocks. Also, under normal operation, a negative mine pressure gradient exists with an exhausting circuit. When fans go down, barometric pressure in the mine rises which temporarily helps keeps methane in coalmines from flowing out of gob areas. This effect is well documented (Kennedy, 1996, p. 221-223)

Specifying the Circuit Fans

Primary fans are either centrifugal or axial. South African mines typically employ large centrifugals while most U.S. and Canadian mines use axials. Both types have advantages. Efficiency, up to 90%, is about the same with either type. Centrifugals are arguably heavier duty, quieter, do not have a pronounced stall region, and can generate higher static pressures (over 30″). Axials are more compact, and airflows can be easily adjusted via blade angle changes. Primary fans range in size from 100 to over 3500 horsepower each. Surface installations containing multiple fans are common for large airflows, and for back-up operation in case one fan goes down for maintenance. Circuit fans can also be installed underground, especially in metal mines.
Primary fans are specified while the circuit is designed. The ventilation engineer must often go back and forth between airway considerations (sizes and numbers), and fan specifications. A fan’s operating point must be located in an efficient part of its curve. Care must be taken to anticipate future circuit changes as mining operations advance.
Fan speed, quantity, pressure, and power are related in equations called the fan laws, which are based on proportionality. The mine ventilation engineer must become familiar with these laws, as described in the 2000 ASHRAE HVAC Systems and Equipment Handbook, pages 18.4-18.5.
The fan installation must be designed after primary fans are selected. The consequences of fan downtime must be carefully considered. This is especially important for coalmines, since methane concentrations can increase when circuit airflows go down. Fans exhausting a mine are typically mounted horizontally near a vertical shaft or borehole. A 90° transition turns the air into the fan inlet. An isolation door is installed between the transition and fan. Coalmines require a blast door to dampen a shock wave caused by a possible methane or coal dust explosion. An evase, or diffuser, is attached to the fan outlet to recover a portion of the velocity pressure exiting the mine. A silencer can be added if surface noise reduction is desired. More and more, variable-speed drives are used with the electric motors to turn the fans. These drives provide soft-start, and variable speed from 50% to 100% of synchronous speed, and even to 110% for temporary emergency duty. The fan installation must be designed for accessibility and ease of maintenance.

Determining Auxiliary System Requirements

Auxiliary fan and duct systems deliver air into dead-end headings. These systems are generally not permitted in coalmines but are common in metal mines. A blowing system is most often used. A fan is set in a fresh air base at the start of a drift, and duct is installed as the drift advances. For drifts under 1000 feet, flexible brattice cloth duct can be used. For longer drifts requiring booster fans, rigid duct is needed since duct gage pressure can drop below atmospheric. Rigid duct also offers less resistance than brattice cloth duct (but it is about eight times more expensive).
The quantity needed at the face is determined by the equipment used and by the rate at which blasting and diesel fumes must be removed. Duct size is selected by the air quantity needed, and by space limitations in the drift. Fans are selected to provide the specified airflow. In general, a single-stage axial fan can generate up to 10″ static water column, which should deliver the required airflow up to 2500 feet through properly sized duct. A larger duct or a two-stage axial fan is needed if distance is much longer. For very long drifts, booster fans are needed about every 2500 feet.
An exhausting system is often used for drifts requiring quick ingress after blasting. Air flows to the face through the drift, captures fumes at the face, and is blown back to the circuit through duct. This keeps the drift clear of fumes. The disadvantages of exhausting systems are that the air picks up heat and humidity en route to the face, rigid duct is required, and the face is not swept by air as with a blowing system. A face overlap fan and duct can be installed.
The 2001 ASHRAE Fundamentals Handbook, page 34.9, offers a friction chart for round duct. However, the ventilation engineer should acquire the friction chart of the specific duct being considered from the supplier. Shock losses through couplings and bends must be tallied. One of the most important considerations is leakage through couplings. This can be minimized by taking special care during installation, by keeping duct pressures under 10” static, and by installing longer pieces of duct (up to 100 feet for brattice cloth, or 20 feet for rigid). Some coupling methods are inherently better than others.
Duct damage is common in mines. Mobile equipment and fly-rock from blasting can punch holes in the duct. These factors can drastically reduce the airflow out of the end. Care must be taken to minimize damage, and to quickly repair or replace damaged pieces.

Assessing the Health and Safety Aspects

Few aspects of underground mining have as direct an impact on the health, safety and morale of the work force as does ventilation. No component of ventilation design should be undertaken without a rigorous review of health and safety aspects. These aspects include:

• Fire and explosion risk.
• Dangerous and toxic substances risk.
• Heat risk.
• Ventilation equipment risk.

For metal mines, fire is the most significant potential ventilation hazard. The ventilation engineer should become familiar with fire science. Fuel, heat, and oxygen are required for combustion. Removing any of these components will prevent combustion. Fuel sources such as oil, diesel fuel and blasting agents are kept in special areas designed to keep out ignition sources. Sprinkler or chemical suppression systems can be installed in these areas as well as in repair shops. Mobile equipment fires are a special concern for modern mining. Vehicles should be fitted with a dry chemical fire suppression system, which is either triggered automatically or by the operator. Electric substation and conveyor fires can be very dangerous.
Spontaneous combustion has plagued both metal and coal mines since antiquity. Fortunately for modern mining, timber is seldom used for ground support although many older mines have worked out areas that contain timber. Coal, being combustible, can be particularly troublesome. Circuits must be designed so that spontaneous combustion fires will not contaminate active workings.
The ventilation engineer should play the “what if” game with ventilation circuits. What are the fire risks associated with any given area? Imagining a fire breaking out in any location. How will the circuits respond? Will fire-induced natural drafts change airflow quantities and directions? How will the fire be detected, how will miners be notified, how will they escape, and how will the fire be fought? MSHA requires that refuge chambers be constructed if miners cannot be hoisted to the surface within one hour of notification. An actual emergency is no time to start thinking of these issues. The ventilation staff must work closely with the Safety Department and mine management in pre-planning how to respond to different emergencies.
For coalmines, methane and coal dust explosions pose the greatest risk. Equipment must be rated “permissible”, or non-sparking. Airways are coated with rock dust to prevent a methane ignition from propagating. Methane is explosive in air from 5% to 15% concentration. Whenever methane reaches 0.25%, MSHA requires that changes be made to improve ventilation. At 0.5%, further steps must be taken, and no other work is permitted until the concentration drops below 0.5%. At 1%, all personnel except those working on ventilation must be vacated from the affected area.
A ventilation system is the first line of defense against toxic or asphyxiating gases. These can be generated by blasting (CO, CO2, NH3, NOX) and by diesel engines (CO, NOX, SO2, various hydrocarbon compounds, and soot). The rock itself can release CO2 and H2S.
The relationship between heat stress and accident frequency has been clearly established in South African mines (Stewart 1982). Work area temperatures should be kept under 85°F wet-bulb, especially where heavy physical work is performed.
Ventilation and air-conditioning equipment may also pose certain health and safety risks. All fan inlets require screens. Fans should be equipped with vibration sensors that can trip the fan if necessary. Silencers may be needed if personnel work in the vicinity. Refrigeration rooms must be well ventilated in case of a sudden refrigerant release. Duct, pipe insulation and other substances such as foam for seals shall be approved by MSHA in accordance with 30 CFR Part 7. Electrical systems must meet rigorous MSHA codes.

A Word of Caution

The discussions presented here are but a very brief overview of the principles used in mine ventilation planning. The person responsible for such planning should either be an experienced engineer, or should work under the direct supervision of such an engineer. The health and safety of the work force are too important to be trusted to inexperienced personnel. Seven English language texts have been written on mine ventilation since 1980 (Hall 1981, Mine Ventilation Society of South Africa 1982, Bossard 1982, McPherson 1993, Kennedy 1996, Hartman 1997 and Tien 1999). The ventilation engineer is strongly encouraged to study these references.

 

REFERENCES

Abbeon Cal, Inc., 2nd ed. 2001. Pocket Ref. T.J. Glover, ed. Sequoia Publishing Inc., Littleton,
CO. p. 213.
Bluhm, S.J. 1981. “Heat Transfer Characteristics of Direct-Contact Crossflow Heat Exchangers”.
Master’s Thesis presented to the University of the Witwatersrand. Republic of South Africa.
Bossard, F. 1982. Chapter 26 – Sources of Heat Entering Mine Air. Mine Ventilation Design
Practices, first ed. Butte, MT, pp. 26-2.
Burrows, J. 1982. Chapter 23 – “Refrigeration – Theory and Operation”. Environmental
Engineering in South African Mines. The Mine Ventilation Society of South Africa.
pp. 631-637.
Duckworth, I. and Mousset-Jones, P. 1993. “A Detailed Heat Balance Study of a Deep Level
Mechanized Cut-and-Fill Stope”, Proceeding of the 6th U.S. Mine Ventilation Symposium,
Bhaskar, R. ed., SME, Inc. Littleton, CO, pp. 441-447.
Fenton, J.L. 1972. Survey of Underground Mine Heat Sources. Master’s Thesis, Montana
College of Mineral Science and Technology, Butte, MT.
Goch, D.C. and Patterson, H.S. 1940. “The Heat Flow into Tunnels.” Journal of the Chemical,
Metallurgical and Mining Society of South Africa, Vol. 41.
Marks, J.R. and Shaffner, L.M. 1993. “An Empirical Analysis of Ventilation Requirements for
Deep Mechanized Stoping at the Homestake Gold Mine”, Proceeding of the 6th U.S. Mine
Ventilation Symposium, Bhaskar, R. ed., SME, Inc. Littleton, CO, pp. 381-385.
Hall, A.E. et.al. 1989. “The Use of Controlled Recirculation to Reduce Heating Costs in
Canada”. Proceedings of the Fourth International Mine Ventilation Congress, S. Gilles, ed.
Pub. by the Australasia Institute of Mining and Metallurgy, Melbourne, Australia.
Hall, CJ. 1981. Mine Ventilation Engineering. Published by the Society of Mining Engineers,
by Lucas-Guinn Co., Inc., Hoboken, New York.
Hartman, H. et.al. ed. 1997. Mine Ventilation and Air Conditioning, 3rd ed., John Wiley & Sons,
Inc., NY. pp. 247-252.
Hartman, H, et. al. 1997. “Heat Indexes and Standards”, Mine Ventilation and Air Conditioning,
3rd ed., John Wiley & Sons, Inc., NY, pp. 610-616.
Howes, M.J., and Nixon, C.A. 1997. “Development of Procedure for Safe Working in Hot
Conditions”, Proceedings of the 6th International Mine Ventilation Congress, Ramani R.V. ed.,
SME, Inc., Littleton, CO pp. 191-197.
Kennedy, W.R. 1996. Practical Mine Ventilation. Intertec Publishing, Chicago, Il., pp 221-223.
McPherson, M.J. 1993. Subsurface Ventilation and Environmental Engineering, Chapman &
Hall, New York.
Patterson, A.M. ed., 1992. Chapters on Planning and Heat transfer. The Mine Ventilation
Practitioner’s Data Book, The Mine Ventilation Society of South Africa.
Patterson, A.M. ed., 1992. Chapter on Planning. The Mine Ventilation Practitioner’s Data Book,
The Mine Ventilation Society of South Africa. pp. P-R 4 to P-R 11.
Sheer, TJ., Butterworth, MD., and Ramsden, R. 2001. Chapter 52. “Ice as a Coolant for Deep
Mines”, Seventh International Mine Ventilation Congress. Published by the Research and
Development Center for Electrical Engineering and Automation in Mining, S. Wasilewski, ed.,
Cracow, Poland.
Stachulak, J. 1989. “Ventilation Strategy and Unique Air Conditioning at INCO Limited”.
Proceedings of the Fourth US Mine Ventilation Symposium. Pub. by the Society for Mining
Metallurgy and Exploration, Berkeley, CA.
Stewart, J.M. Chapter 21, “Practical Aspects of Human Heat Stress”. Environmental
Engineering in South African Mines, Burrows et. al. ed., The Mine Ventilation Society of
South Africa, pp. 574.
Tien, J.C. 1999. Chapter 20 – “Controlled Recirculation”, Practical Mine Ventilation
Engineering, Intertec Publishing Corp. Chicago, Il. pp. 370-392.
van der Walt, J. and van Rensburg, CSJ. 1988. “The Options for Cooling Deep Mines”, Fourth
International Mine Ventilation Congress. Australasian Institute of Mining and Metallurgy,
Melborne, Australia. pp. 401.
Whillier, A. 1977. “Predicting the Performance of Forced-Draught Cooling Towers”, Journal of
the Mine Ventilation Society of South Africa, Vol. 30, No. 1, pp 2-25.
Whillier and Thorpe; reported by Hemp, R. Chapter 22: “Sources of Heat in Mines”,
Environmental Engineering in South African Mines, Burrows et. al. ed., The Mine Ventilation
Society of South Africa, pp. 574.

Thanks- PMJ