# Rules of Thumb for Mining and processing



## alshangiti (13 مايو 2007)

*Rule Number TopicRule of Thumb1.01DiscoveryIt takes 25,000 claims staked to find 500 worth diamond drilling to find one mine. Source: Lorne Ames1.02DiscoveryOn average, the time between discovery and actual start of construction of a base metal mine is 10 years; it is less for a precious metal mine. Source: J.P. Albers1.03DiscoveryOn average, the time between discovery and actual start of production of a mine in an established mining district (“brown field”) is seven years. Source: Sylvain Paradis1.04DiscoveryOn average, the time between discovery and actual start of production of a mine in a district where there is no previously established mining activity (“green field”) is ten years. Source: Sylvain Paradis1.05CostsThe amount expended on diamond drilling and exploration development for the purposes of measuring a mineral resource should approximately equal 2% of the gross value of the metals in the deposit. Source: Joe Gerden1.06Bulk SampleThe minimum size of a bulk sample, when required for a proposed major open pit mine is in the order of 50,000 tons (with a pilot mill on site). For a proposed underground mine, it is typically only 5,000 tons. Source: Jack de la Vergne1.07Ore Reserve EstimateThe value reported for the specific gravity (SG) of an ore sample on a metallurgical test report is approximately 20% higher than the correct value to be employed in the resource tonnage calculation. Source: Jack de la Vergne1.08Ore Resource EstimateTo determine an “inferred” or “possible” resource, it is practice to assume that the ore will extend to a distance at least equal to half the strike length at the bottom of measured reserves. Another rule is that the largest horizontal cross section of an ore body is half way between its top and bottom. Source: H. E. McKinstry1.09Ore Resource EstimateIn the base metal mines of Peru and the Canadian Shield, often a zonal mineralogy is found indicating depth. At the top of the ore body sphalerite and galena predominate. Near mid-depth, chalcopyrite becomes significant and pyrite appears. At the bottom, pyrite, and magnetite displace the ore. Source: H. E. McKinstry1.10Ore Resource EstimateArchean aged quartz veins are generally two times as long as their depth extent, but gold zones within these vein systems are 1/5 - 1/10 as long as their depth extent. Source: Gord Yule1.11Ore Resource EstimateIn gold mines, the amount of silver that accompanies the gold may be an indicator of depth. Shallow gold deposits usually have relatively high silver ******* while those that run deep have hardly any. Source: James B. Redpath1.12Ore Resource EstimateAs a rule of thumb, I use that 2P (Probable) reserves are only such when drill spacing does not exceed five to seven smallest mining units (SMU). Open pit mining on 15m benches could have an SMU of 15m by 15m by 15m. Underground, an SMU would be say 3m by 3m by 3m (a drift round). Source: René Marion1.13Ore Resource EstimateYour thumb pressed on a 200-scale map covers 100,000 tons of ore per bench (height assumed to be 50 feet). Source: Janet Flinn1.14Strike and DipThe convention for establishing strike and dip is always the Right Hand Rule. With right hand palm up, open and extended, point the thumb in the down-dip direction and the fingertips provide the strike direction. Source: Mike Neumann*


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## ingmauritania (15 مايو 2007)

good afternoon


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## alshangiti (16 مايو 2007)

*Rule Number 2 -Rule of Thumb for Rock mechanics*

.01Ground StressThe vertical stress may be calculated on the basis of depth of overburden with an accuracy of ± 20%. This is sufficient for engineering purposes. Source: Z.T. Bieniawski2.02Ground StressDiscs occur in the core of diamond drill holes when the radial ground stresses are in excess of half the compressive rock strength. Source: Obert and Stephenson2.03Ground StressThe width of the zone of relaxed stress around a circular shaft that is sunk by a drill and blast method is approximately equal to one-third the radius of the shaft excavation. Source: J. F. Abel2.04Ground ControlThe length of a rock bolt should be one-half to one-third the heading width. Mont Blanc Tunnel Rule (c.1965)2.05Ground ControlIn hard rock mining, the ratio of bolt length to pattern spacing is normally 1½:1. In fractured rock, it should be at least 2:1. (In civil tunnels and coalmines, it is typically 2:1.) Source: Lang and Bischoff (1982)2.06Ground ControlIn mining, the bolt length/bolt spacing ratio is acceptable between 1.2:1 and 1.5:1. Source: Z.T. Bieniawski (1992)2.07Ground ControlIn good ground, the length of a roof bolt can be one-third of the span. The length of a wall bolt can be one-fifth of the wall height. The pattern spacing may be obtained by dividing the rock bolt length by one and one-half. Source: Mike Gray (1999)2.08Ground ControlThe tension developed in a mechanical rock bolt is increased by approximately 40 Lbs. for each one foot-pound increment of torque applied to it. Source: Lewis and Clarke2.09Ground ControlA mechanical rock bolt installed at 30 degrees off the perpendicular may provide only 25% of the tension produced by a bolt equally torqued that is perpendicular to the rock face, unless a spherical washer is employed. Source: MAPAO2.10Ground ControlFor each foot of friction bolt (split-set) installed, there is 1 ton of anchorage. Source: MAPAO2.11Ground ControlThe shear strength (dowel strength) of a rock bolt may be assumed equal to one-half its tensile strength. Source: P. M. Dight2.12Ground ControlThe thickness of the beam (zone of uniform compression) in the back of a bolted heading is approximately equal to the rock bolt length minus the spacing between them. Source: T.A. Lang2.13Ground ControlHoles drilled for resin bolts should be ¼ inch larger in diameter than the bolt. If it is increased to 3/8 inch, the pull out load is not affected but the stiffness of the bolt/resin assembly is lowered by more than 80%, besides wasting money on unnecessary resin. Source: Dr. Pierre Choquette2.14Ground ControlHoles drilled for cement-grouted bolts should be ½ to 1 inch larger in diameter than the bolt. The larger gap is especially desired in weak ground to increase the bonding area. Source: Dr. Pierre Choquette2.15Ground ControlEvery 100° F rise in temperature decreases the set time of shotcrete by 1/3. Source: Baz-Dresch and Sherril2.16Mine DevelopmentPermanent underground excavations should be designed to be in a state of compression. A minimum safety factor (SF) of 2 is generally recommended for them. Source: Obert and Duval2.17Mine DevelopmentThe required height of a rock pentice to be used for shaft deepening is equal to the shaft width or diameter plus an allowance of five feet. Source: Jim Redpath2.18Stope Pillar and DesignA minimum SF of between 1.2 and 1.5 is typically employed for the design of rigid stope pillars in hard rock mines. Various Sources2.19Stope Pillar and DesignFor purposes of pillar design in hard rock, the uniaxial compressive strength obtained from core samples should be reduced by 20-25% to obtain a true value underground. The reduced value should be used when calculating pillar strength from formulas relating it to compressive strength, pillar height, and width (i.e. Obert Duval and Hedley formulas). Source: C. L. de Jongh2.20Stope Pillar and DesignThe compressive strength of a stope pillar is increased when later firmly confined by backfill because a triaxial condition is created in which s3 is increased 4 to 5 times (by Mohr’s strength theory). Source: Donald Coates2.21SubsidenceIn block caving mines, it is typical that the cave is vertical until sloughing is initiated after which the angle of draw may approach 70 degrees from the horizontal, particularly at the end of a block. Source: Fleshman and Dale2.22SubsidencePreliminary design of a block cave mine should assume a potential subsidence zone of 45-degrees from bottom of the lowest mining level. Although it is unlikely that actual subsidence will extend to this limit, there is a high probability that tension cracking will result in damage to underground structures (such as a shaft) developed within this zone. Source: Scott McIntosh2.23SubsidenceIn hard rock mines employing backfill, any subsidence that may occur is always vertical and nothing will promote side sloughing of the cave (even drill and blast). Source: Jack de la Vergne2.24Rockbursts75% of rockbursts occur within 45 minutes after blasting. Source: Swanson and Sines2.25RockburstsThe larger the rockburst, the more random the pattern in time of occurrence. Microseismic data from many areas shows that the smaller microseismic events tend to be concentrated at or just after blast time, on average (see above). However, the larger the event, the more random its time of occurrence. Source: Richard Brummer2.26RockburstsIn burst prone ground, top sills are advanced simultaneously in a chevron (‘V’) pattern. Outboard sills are advanced in the stress shadow of the leading sill with a lag distance of 24 feet. Source: Luc Beauchamp2.27RockburstsSeismic events may be the result of the reactivation of old faults by a new stress regime. By Mohr-Coulomb analysis, faults dipping at 30 degrees are the most susceptible; near vertical faults are the safest. Source: Asmis and Lee2.28RockburstsThere can be little doubt that it is possible to control violent rock behavior by means of preconditioning or de-stressing under appropriate circumstances. This technology, therefore, has the potential to be profitably harnessed for use in the mining of deeper orebodies, particularly hazardous situations such as highly stressed high grade remnants, or development into areas known to be prone to bursting. Source: Board, Blake & Brummer


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## هانى شرف الدين (17 مايو 2007)

موضوعاتك رائعة جزاك الله خيرا


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## alshangiti (21 مايو 2007)

*Rule Number 3 -Rule of Thumb for Mining Methods*

3.01 Method SelectionA flatly dipping ore body may be mined using Blasthole when the height of ore exceeds 100 feet (30m); otherwise, it is mined Room and Pillar. Source: John Folinsbee
3.02 InclinationOre will not run on a footwall inclined at less than 50 degrees from the horizontal. Source: Fred Nabb
3.03InclinationEven a steeply dipping ore body may not be drawn clean of ore by gravity alone. A significant portion of the broken ore will inevitably remain (“hang”) on the footwall. If the dip is less than 60 degrees, footwall draw points will reduce, but not eliminate, this loss of ore. Source: Chen and Boshkov
3.04Stope DevelopmentThe number of stopes developed should normally be such that the planned daily tonnage can be met with 60% to 80% of the stopes. The spare stopes are required in the event of an unexpected occurrence and may be required to maintain uniform grades of ore to the mill. This allowance may not be practical when shrinkage is applied to a sulfide ore body, due to oxidation. Source: Folinsbee and Nabb
3.05Stope DevelopmentIn any mine employing backfill, there must be 35% more stoping units than is theoretically required to meet the daily call (planned daily tonnage). Source: Derrick May
3.06Ore WidthBlasthole (longhole) Stoping may be employed for ore widths as narrow as 3m (10 feet). However, this narrow a width is only practical when there is an exceptionally good contact separation and a very uniform dip. Source: Clarke and Nabb
3.07Ore WidthSequence problems are not likely in the case of a massive deposit to be caved if the horizontal axes are more than twice the proposed draw height. Source: Dennis Laubscher
3.08Footwall DriftsFootwall drifts for blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth, it should be not less than 30m (100 feet). Source: Jack de la Vergne
3.09DilutionA ton of ore left behind in a stope costs you twice as much as milling a ton of waste rock (from dilution). Source: Peter J. George


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## alshangiti (23 مايو 2007)

*Rule of thumb no- 3 -Mine layout*

4.01 Pit LayoutThe overall slope (including berms, access roads, and haul roads) of large open pits in good ground will eventually approach the natural angle of repose of broken wall rock (i.e. 38 degrees), except for the last few cuts, which may be steeper. Source: Jack de la Vergne 

4.02 Pit LayoutWhen hard laterites are mined in an open pit, safe pit slopes may be steeper than calculated by conventional practice (as steep as 50 degrees between haul roads). Source: Companhia Vale do Rio Doce

4.03Pit LayoutFor haul roads in general, 10% is the maximum safe sustained grade. For particular conditions found at larger operations, the grade has often been determined at 8%. It is usually safe to exceed the maximum sustained grade over a short distance. Source: USBM4.04
Pit LayoutThe maximum safe grade over a short distance is generally accepted to be 15%. It may be 12% at larger operations. Source: Kaufman and Ault4.05Pit LayoutThe maximum safe operating speed on a downhill grade is decreased by 2 km/h for each 1% increase in gradient. Source: Jack de la Vergne4.06Pit LayoutEach lane of travel should be wide enough to provide clearance left and right of the widest haul truck in use equal to half the width of the vehicle. For single lane traffic (one-way), the travel portion of the haul road is twice the width of the design vehicle. For double lane (two-way), the width of roadway required is 3½ times the width of the widest vehicle. Source: Association of American State Highway Officials (AASHO)4.07Pit LayoutTo avoid a collision caused by spinout, the width of an open pit haul road should equal the width plus the length of the largest truck plus 15 feet safety distance. Source: Janet Flinn4.08Pit LayoutA crushed rock safety berm on a haulage road should be at least as high as the rolling radius of the vehicle tire. A boulder-faced berm should be of height approximately equal to the height of the tire of the haulage vehicle. Source: Kaufman and Ault4.09Crown PillarA crown pillar of ore beneath the open pit is usually left in place while underground mining proceeds. The height of the crown pillar in good ground is typically made equal to the maximum width of stopes to be mined immediately beneath. When the overburden is too deep, the ore body is not mined by open pit, but a crown pillar is left in place of height the same as if it were. If the outcrop of the ore body is badly weathered (“oxidized”) or the ore body is cut by major faults, under a body of water or a muskeg swamp - the height of the crown pillar is increased to account for the increased risk. Source: Ron Haflidson and others4.10Mine EntriesSmall sized deposits may be most economically served by ramp and truck haulage to a vertical depth of as much as 500m (1,600 feet). Source: Ernie Yuskiw4.11Mine EntriesA medium-sized deposit, say 4 million (short) tons, may be most economically served by ramp and truck haulage to a vertical depth of 250m (800 feet). Source: Ernie Yuskiw4.12Mine EntriesThe optimum “changeover” depth from ramp haulage to shaft hoisting is 350m (1,150 feet). Source: Northcote and Barnes4.13Mine EntriesIn good ground, at production rates less than one million tons per year, truck haulage on a decline (ramp) is a viable alternative to shaft hoisting to depths of at least 300m. Source: G.G. Northcote4.14Mine EntriesWestern Australia practice suggests a depth of 500m or more may be the appropriate transition depth from decline (ramp) haulage to shaft hoisting. Source: McCarthy and Livingstone 4.15Mine EntriesProduction rates at operating mines were found to range from 38% to 89% of the estimated truck fleet capacity. For a proposed operation, 70% is considered to be a reasonable factor for adjusting theoretical estimates to allow for operating constraints. Source: McCarthy and Livingstone4.16Mine EntriesShallow ore bodies mined at over 5,000 tpd are more economically served by belt conveyor transport in a decline entry than haul trucks in a ramp entry. Source: Al Fernie4.17Mine EntriesAs a rule, a belt conveyor operation is more economical than rail or truck transport when the conveying distance exceeds one kilometer (3,281 feet). Source: Heinz Altoff4.18ShaftsThe normal location of the production shaft is near the center of gravity of the shape (in plan view) of the ore body, but offset by 200 feet or more. Source: Alan O’Hara4.19ShaftsThe first lift for a near vertical ore body should be approximately 2,000 feet. If the ore body outcrops, the shaft will then be approximately 2,500 feet deep to allow for gravity feed and crown pillar. If the outcrop is or is planned to be open cut, the measurement should be made from the top of the crown pillar. If the ore body does not outcrop, the measurement is taken from its apex. Source: Ron Haflidson4.20ShaftsThe depth of shaft should allow access to 1,800 days mining of ore reserves. Source: Alan O’Hara4.21ShaftsFor a deep ore body, the production and ventilation shafts are sunk simultaneously and positioned within 100m or so of each other. Source: D.F.H. Graves4.22Underground LayoutFootwall drifts for blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth it should be not less than 30m (100 feet). Source: Jack de la Vergne4.23Underground LayoutOre passes should be spaced at intervals not exceeding 500 feet (and waste passes not more than 750 feet) along the footwall drift, when using LHD extraction. Source: Jack de la Vergne4.24Underground LayoutThe maximum economical tramming distance for a 5 cubic yard capacity LHD is 500 feet, for an 8 cubic yard LHD it is 800 feet. Source: Len Kitchener4.25Underground LayoutThe amount of pre-production stope development required to bring a mine into production is equal to that required for 125 days of mining. Source: Alan O’Hara


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## alshangiti (28 مايو 2007)

*Rule Number 4 -Rule of Thumb for Enviromental*

*Environmental Impact Statement The cost of an environmental impact statement (EIS) (including base line monitoring and specific previously performed studies) may cost approximately 2.5% of the total pre-production capital cost for a plain vanilla domestic mining project. The cost can increase by 2% for an undertaking that is politically or environmentally sensitive. In the latter case, the cost may increase further if proposals are challenged in the courts. Source: R.W. Corkery *


*5.02 Site Layout If the mill (concentrator) is located close to the mine head, the environmental impact is reduced and so are the costs. Pumping tailings from the mill is cleaner, less disruptive to the terrain, and less expensive than to truck haul ore over a similar distance. When pumping water to the mill and hauling concentrate from the mill is considered, the argument is usually stronger. The rule is further reinforced in the case of an underground mine where a portion of the tailings is dedicated for paste fill or hydraulic fill. Source: Edgar Köster *


*5.03 Site Layout The mine administration offices should be located as near as possible to the mine head to reduce the area of disturbance, improve communications, and reduce transit time. Source: Brian Calver *


*5.04 Site Layout When a mine has a camp incorporated into its infrastructure, the campsite should be as close as practical to the mine to minimize the impact from service and utility lines, decrease the area of the footprint of disturbance, shorten travel time, and reduce costs. Source: George Greer *


*5.05 Site Drainage and Spill Protection Drainage ditches to protect the mine plant should be designed to develop peak flow rates based on 100 year, 24 hour storm charts. Source: AASHO *
*5.06 Site Drainage and Spill Protection Dykes around tank farms should be designed to hold 100% of the capacity of the largest tank + 10% of the capacity of the remaining tanks. Source: George Greer *


*5.07 Water Supply If a drilled well is to be used for fire fighting without additional storage, it should demonstrate (by pumping test) a minimum capacity of 40 USGPM continuously for two hours during the driest period of the year. Various Sources *
*5.08 Water Supply Chlorine should be added to water at a rate of approximately 2 mg/litre to render it safe to drink. Source: Ontario Ministry of Health and Welfare *


*5.09 Dust Suppression Dust emissions emanating from the transport of ore will not remain airborne when the size of dust particle exceeds 10 m (ten microns). Source: Howard Goodfellow*


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## alshangiti (3 يوليو 2007)

*Rule Number 5 -Rule of Thumb for mine maintenace*

*The degree of maintenance enforcement at an operating mine should be just less than the point that disruptions to operations are at a level where additional maintenance costs equal the resulting profits from production. Source: David Chick 

28.02 General In a trackless mine operating round the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. Source: John Gilbert 
28.03 General Emergency repairs should not exceed 15% of the maintenance workload. Source: John Rushton 

28.04 General LHD units at a shallow mine with ramp entry should have a utilization of 5,000 - 6,000 hours per year. Source: Unknown 
28.05 General Captive LHD units should have a utilization of 3,500 - 4,500 hours per year. Source: Unknown 

28.06 General LHD units in production service should have a useful life of at least 12,000 hours, including one rebuild at 7,500 hours. A longer life can be presumed from LHD units at the high end of the market with on-board diagnostics. Source: John Gilbert 

28.07 General Underground haul trucks should have a useful life of 20,000 hours; more if they are electric (trolley system). Longer life may be presumed in the light of today’s improved onboard diagnostics and better management of equipment maintenance in general. Source: John Chadwick 
28.08 Service An efficient Maintenance Department should be able to install one dollar worth of parts and materials for less than one dollar of labor cost. Source: John Rushton 

28.09 Service A servicing accuracy of 10% is a reasonable goal. In other words, no unit of equipment should receive the 250-hour service at more than 275 hours. Source: Larry Widdifield 

28.10 Infrastructure With ramp entry, a satellite shop is required when the mean mining depth reaches 200m below surface. A second one is required at a vertical depth of 400m. Source: Jack de la Vergne 
28.11 Infrastructure With ramp and shaft entry, a main shop is required underground when the mean mining depth reaches 500m below surface. Source: Jack de la Vergne 

28.12 Infrastructure A main shop facility underground should have the capacity to handle 10% of the underground fleet. Source: Keith Vaananen 
28.13 Infrastructure Service shops for open pit mines should be designed with plenty of room between service bays for lay-down area. As a rule of thumb, the width of the lay-down between bays should be at least equal to the width of the box of a pit truck. Source: Cass Atkinson 

28.14 Infrastructure Surface shops should be designed with one maintenance bay for six haul trucks having a capacity of up to 150 tons. This ratio is 4:1 for larger trucks. The shops should also include one tire bay and two lube bays. Additional maintenance bays are required for service trucks (1:20) and support equipment (1:12). Source: Don Myntii*​


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## alshangiti (23 يوليو 2007)

*rule of thumb mineral economomics*

Metal PriceThe long-term average price of a common mineral commodity (the price best used for economic evaluation in a feasibility study) is 1.5 times the average cost of production, worldwide. Source: Sir Ronald Prain

Pre-production Capital CostThe pre-production capital cost estimate (Capex) should include all construction and operating expenses until the mine has reached full production capacity or three months after reaching 50% of full capacity, whichever occurs first. This is the basic transition point between capital and operating costs. Source: John Halls 

 expenditure includes all costs of construction and mine development until three months after the mine has reached 25% of its rated production capacity. Source: Jon Gill7.04Cash FlowThe total cash flow must be sufficient to repay the capital cost at least twice. Source: L. D. Smith

Cash FlowProject loans should be repaid before half the known reserves are consumed. Source: G.R Castle

Cash FlowIncremented cash flow projections should each be at least 150% of the loan repayment scheduled for the same period. Source: G.R. Castle

Cash FlowThe operating cost should not exceed half the market value of minerals recovered. Source: Alan Provost7.08Net Present ValueThe discount factor employed to determine the NPV is often 10%; however, it should be Prime + 5%. Source: G.R. Castle

Net Present ValueThe increment for risk may add 4% to 6% to the base opportunity cost of capital in the determination of a discount rate. Source: Bruce Cavender

Net Present ValueThe value of the long-term, real (no inflation) interest rate is 2.5%. This value is supported by numerous references in the literature. Source: L.D. Smith

Net Present ValueIn numerous conversations with managers of mining firms, I have found that 15% in real terms is the common discount rate used for decision purposes. Source: Herbert Drecshler

 (1980)Net Present ValueIn 1985, the discount rates of many 
mining companies raged from 14% to15%. Source: H. J.

Net Present ValueThe true present value (market value) of a project determined for purposes of joint venture or outright purchase is equal to half the NPV typically calculated. Source: J. B.
 Redpath

Rate of ReturnThe feasibility study for a hard rock mine should demonstrate an internal rate of return (IRR) of at least 20% – more during periods of high inflation. Source: J. B. Redpath

Working CapitalWorking capital equals ten weeks operating cost plus cost of capital spares and parts. Source: Alan O’Hara

Working CapitalWorking capital is typically ten weeks of operating cost plus the spare parts inventory. Source: METSInfo

Closure Costs The salvage value of plant and equipment should pay for the mine closure costs. Source: Ron Haflidson7.18Closure Costs For purposes of cash flow, the cost of reclamation used to be equated with the salvage value of the mine plant, but this is no longer valid in industrialized nations. Source: Paul Bartos


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## محمد الطاهير (8 مارس 2008)

I NEED HELP,
if i can recieve any books about economics and investment in exploration and production upstream and how to develop and produce a gas or oil field: number of appraisal wells, kind of surface equipment.....


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## alshangiti (28 سبتمبر 2011)

maintenance 

*Mining Field: *Mine Maintenance 
*Area: *General •The degree of maintenance enforcement at an operating mine should be just less than the point that disruptions to operations are at a level where additional maintenance costs equal the resulting profits from production. _Source: David Chick_ •In a trackless mine operating round the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. _Source: John Gilbert_ •Emergency repairs should not exceed 15% of the maintenance workload. _Source: John Rushton_ •LHD units at a shallow mine with ramp entry should have a utilization of 5,000 - 6,000 hours per year. _Source: Unknown_ •Captive LHD units should have a utilization of 3,500 - 4,500 hours per year. _Source: Unknown_ •LHD units in production service should have a useful life of at least 12,000 hours, including one rebuild at 7,500 hours. A longer life can be presumed from LHD units at the high end of the market with on-board diagnostics. _Source: John Gilbert_ •Underground haul trucks should have a useful life of 20,000 hours; more if they are electric (trolley system). _Source: John Chadwick_ 
*Area: *Service •An efficient Maintenance Department should be able to install one dollar worth of parts and materials for less than one dollar of labor cost. _Source: John Rushton_ •A servicing accuracy of 10% is a reasonable goal. In other words, no unit of equipment should receive the 250-hour service at more than 275 hours. _Source: Larry Widdifield_ 
*Area: *Infrastructure •With ramp entry, a satellite shop is required when the mean mining depth reaches 200m below surface. A second one is required at a vertical depth of 400m. _Source: Jack de la Vergne_ •With ramp and shaft entry, a main shop is required underground when the mean mining depth reaches 500m below surface. _Source: Jack de la Vergne_ •A main shop facility underground should have the capacity to handle 10% of the underground fleet. _Source: Keith Vaananen_ •Service shops for open pit mines should be designed with plenty of room between service bays for lay-down area. As a rule of thumb, the width of the lay-down between bays should be at least equal to the width of the box of a pit truck. _Source: Cass Atkinson_


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## alshangiti (28 سبتمبر 2011)

*Mining Field: *Explosives and Drilling 
*Area: *Powder Consumption •Powder consumption for shaft sinking is 2.5 lb./short ton broken.
Listed below is typical powder consumption in hard rock. 
Shaft Sinking – 2.5 Lb./short ton broken
Drifting – 1.8 Lb./short ton broken
Raising – 1.5 Lb./short ton broken
Slashing – 0.8 Lb./short ton broken
Shrink Stope – 0.5 Lb./short ton broken
O/H Cut and Fill – 0.5 Lb./short ton broken
Bulk Mining – 0.4 Lb./short ton broken
Block Cave u/c – 0.1 Lb./short ton to be caved
Open Pit Cut – 0.9 Lb./short ton broken
Open Pit Bench – 0.6 Lb./short ton broken
_Source: Various_ 
*Area: *Explosive Choice •The strength of pure ammonium nitrate (AN) is only about one-third as great as that of an oxygen balanced mixture with fuel oil (ANFO). _Source: Dr. Melvin Cook_ 
*Area: *Blasting Strength •Blasting strength is a direct function of density, other things being equal. Typical explosives for dry ground (ANFO) may have a blasthole density (specific gravity) of 0.8 to 1.3, while for wet ground (slurry or emulsion) it varies from 1.1 to 1.3. Developments in explosive technology make it possible to choose any density desired, within the given ranges. _Source: Dr. Nenad Djordjevic_ 
*Area: *Spacing and Burden •For hard rock open pits or backfill rock quarries, the burden between rows can vary from 25 to 40 blasthole diameters. Spacing between holes in a row can vary between 25 and 80 blasthole diameters. _Source: Dr. Nenad Djordjevic_ •To obtain optimum fragmentation and minimum overbreak for hard rock open pits or backfill rock quarries, the burden should be about one-third the depth of holes drilled in the bench. _Source: Dr. Gary Hemphill_ •To obtain optimum fragmentation and minimum overbreak for stripping hard rock open pits or quarrying rock fill, the burden should be about 25 times the bench blasthole diameter for ANFO and about 30 times the blasthole diameter for high explosives. _Source: Dr. Gary Hemphill_ •The burden required in an open pit operation is 25 times the hole diameter for hard rock, and the ratio is 30:1 and 35:1 for medium and soft rock, respectively. The spacing is 1 to 1.5 times the burden and the timing is a minimum of 5 ms (millisecond) per foot of burden. _Source: John Bolger_ •The burden and spacing required in the permafrost zones of the Arctic is 10-15% less than normal. _Source: Dr. Ken Watson_ •When "smooth wall" blasting techniques are employed underground, the accepted standard spacing between the trim (perimeter) holes is 15-16 times the hole diameter and the charge in perimeter holes is 1/3 that of the regular blastholes. The burden between breast holes and trim holes is 1.25 times the spacing between trim holes. _Source: M. Sutherland_ 
*Area: *Collar Stemming •The depth of collar for a blasthole in an open pit or quarry is 0.7 times the burden. _Source: John Bolger_ •The depth of collar stemming is 20-30 times the borehole diameter. _Source: Dr. Nenad Djordjevic_ •For open pits or back-fill rock quarries, pea gravel of a size equal to 1/17 the diameter of the blasthole should be employed for collar stemming (i.e. ½ inch pea gravel for an 8½-inch diameter hole). _Source: Dr. Gary Hemphill_ 
*Area: *Relief Holes •Using a single relief hole in the burn cut, the length of round that can be pulled in a lateral heading is 3 feet for each inch diameter of the relief hole. For example, a 24-foot round can be pulled with an 8-inch diameter relief hole. _Source: Karl-Fredrik Lautman_ •It has been found that a relief hole of 250 mm (10 inches) will provide excellent results for drift rounds up to about 9.1m (30 feet) in length. _Source: Bob Dengler_ 
*Area: *Blastholes •The cost of drilling blastholes underground is about four times the cost of loading and blasting them with ANFO. Present practice is usually based on the historical use of high explosives where the costs were about equal. An opportunity exists for savings in cost and time for lateral headings greater than 12 feet by 12 feet in cross-section by drilling the blastholes to a slightly larger diameter than is customary. _Source: Jack de la Vergne_ •The "subdrill" (over-drill) for blastholes in open pits is 0.3 times the burden in hard rock and 0.2 times the burden in medium/soft rock. _Source: John Bolger_ •"Sub-grade" (over-drill) is in the order of 8 to 12 blasthole diameters. _Source: Dr. Nenad Djordjevic_ 
*Area: *Ground Vibration •The ground vibration produced by the first delay in a burn cut round is up to five times higher than that generated by subsequent delays well away from the cut. _Source: Tim Hagan_ 
*Area: *Crater Blasting •Crater blasting will be initiated if the charge acts as a sphere, which in turn requires the length of a decked charge in the blasthole to be no more than six times its diameter. _Source: Mining Congress Journal_ 
*Area: *Labor Cost •The labor cost for secondary blasting can be expressed as a percentage of the labor cost for primary mucking. For Sub-Level Cave and Crater Blasthole stoping, it is around 30%; for Sub-Level Retreat it is closer to 10%. _Source: Geoff Fong_ 
*Area: *Drilling •Percussion drilling is required for drilling blastholes in rocks with a hardness of 4 or greater on the Mohs' scale (Refer to Chapter 1). These are mainly the volcanic rocks. Rotary drilling is satisfactory for softer rocks, mainly sedimentary. _Source: Dr. Gary Hemphill_


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## alshangiti (2 أكتوبر 2011)

*Mining Field: *Mine Dewatering 
*Area: *Water Balance •The average consumption of service water for an underground mine is estimated at 30 US gallons per ton of ore mined per day. The peak consumption (for which the water supply piping is designed) can be estimated at 100 USGPM per ton of ore mined per day. _Source: Andy Pitz_ •Ore hoisted from an underground hard rock mine has moisture ******* of approximately 3%. _Source: Larry Cooper_ •A water fountain left running underground wastes 1,100 USGPD. _Source: Jack de la Vergne_ •A diesel engine produces 1.2 litres (or gallons) of moisture for each litre (or gallon) of fuel consumed. _Source: John Marks_ •In the hard rock mines of the Canadian Shield, ground water is seldom encountered by mine development below 450m (1,500 feet). This may be because the increased ground stress at depth tends to close the joints and fractures that normally conduct water. _Source: Jim Redpath_ 
*Area: *Layout •The main pump station underground must have sufficient excavations beneath it to protect from the longest power failure. The suggested minimum capacity of the excavations is 24 hours and a typical design value is 36 hours. _Source: Jack de la Vergne_ •The main pumps should be placed close to the sump so that the separation will allow for a minimum straight run of pipe equal to five times (preferably ten times) the diameter of the pipe. _Source: Various_ •Turbulence will be sufficient to ensure good mixing of a flocculating agent if the water velocity is at least 1m/s and maintained for 30 seconds in a feed pipe or channel. _Source: NMERI of South Africa_ 
*Area: *Design •Piping for long runs should be selected on the basis that the water velocity in the pipe will be near 10 feet/sec (3m/s). The speed may be increased up to 50% in short runs. _Source: Various_ •In underground mines, static head is the significant factor for pump design if the pipes are sized properly. To obtain the total head, 5 -10% may be added to the static head to account for all the friction losses without sacrificing accuracy. _Source: Andy Pitz_ •Pump stations for a deep mine served by centrifugal pumps are most economically placed at approximately 2,000-foot (600m) intervals. _Source: Andy Pitz_ •The outlet velocity of a centrifugal pump should be between 10 and 15 feet per second to be economical. _Source: Queen's University_ •Centrifugal pumps should not operate at a speed exceeding 1,800 RPM (except for temporary or small pumps that may operate at 3,600 RPM). This is because impeller wear is proportional to the 2.5 power of the speed. In other words, half the speed means nearly six times the impeller life. _Source: Canadian Mine Journal_ •The maximum lift of a centrifugal pump is a function of the motor torque, which in turn is a function of the supply voltage. Since it is a squared function, a 10% drop in line voltage can result in a 20% loss in head. _Source: Jack de la Vergne_ •The velocity of dirty water being pumped should be greater than 2 fps in vertical piping and 5 fps in horizontal piping. These speeds are recommended to inhibit solids from settling. _Source: GEHO_ •Slime particles less than 5m in diameter cannot be precipitated without use of a flocculating agent. _Source: B. N. Soutar_


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## alshangiti (4 أكتوبر 2011)

*Mining Field: *Crushers and Rockbreakers 
*Area: *Crusher Selection •For a hard rock mine application below 600 tonnes/hour, select a jaw as the primary crusher. Over 1,000 tph, select a gyratory crusher. Between these capacities, you have a choice. _Source: Chris Ottergren_ •For a hard rock mine application below 540 tonnes/hour, a jaw crusher is more economical. Above 725 tonnes/hour, jaw crushers cannot compete with gyratory crushers at normal settings (6 -10 inches). _Source: Lewis, Cobourn and Bhappu_ •For an underground hard rock mine, a gyratory crusher may be more economical in the case where its required daily production exceeds 8,000 tonnes of ore. _Source: Jack de la Vergne_ •If the hourly tonnage to be crushed divided by the square of the required gape in inches is less than 0.115, use a jaw crusher; otherwise use a gyratory. (If the required capacity in metric tph is less than 162 times the square of the gape in metres, use a jaw crusher.) _Source: Arthur Taggart_ •Nearly all crushers produce a product that is 40% finer than one-half the crusher setting. _Source: Babu and Cook_ •The product of a jaw crusher will have a size distribution such that the -80% fraction size (d80) is slightly less than the open-side setting of the crusher. For example, if the open-side setting is 6 inches, the d80 product size will be 5¾ inches. _Source: Unknown_ •In a hard rock mine, the product from a jaw crusher will tend to be slabby, while the product from a gyratory crusher may tend to be blocky, the latter being easier to convey through transfer points on a conveyor system. _Source: Heinz Schober_ •Impact crushers (rotary or hammer mills) have the capacity for high reduction ratios (up to 40:1), but are rarely applied to hard rock mines. Since they depend on high velocities for crushing, wear is greater than for jaw or gyratory crushers. Hence, they should not be used in hard rock mines that normally have ores containing more than 15% silica (or any ores that are abrasive). _Source: Barry Wills_ 
*Area: *Crusher Design •The approximate capacity of a jaw crusher for hard rock application at a typical setting may be obtained by multiplying the width by 10 to get tonnes per hour. For example, a 48 by 60 crusher will have a capacity in the order of 600 tph when crushing ore in a hard rock mine. _Source: Jack de la Vergne_ •The capacity of a jaw crusher selected for underground service should be sufficient to crush the daily requirement in 12 hours. _Source: Dejan Polak_ •For most applications, 7:1 is the maximum practical reduction factor (ratio) for a jaw crusher, but 6:1 represents better design practice. _Source: Jack de la Vergne_ •For most applications, 6:1 is the maximum practical reduction factor (ratio) for a cone crusher, but 5:1 represents better design practice. _Source: Jack de la Vergne_ •Corrugated liner plates designed for jaw crushers (to avoid a slabby product) result in shortening liner life by up to two-thirds and they are more prone to plugging than smooth jaws. _Source: Ron Doyle_ 
*Area: *Crusher Installation •The crushed ore surge pocket beneath a gyratory crusher should have a live load capacity equal to 20 minutes of crusher capacity or the capacity of two pit trucks. _Source: Various_ •It will take six months to excavate, install, and commission an underground crusher station for a typical jaw crusher. For a very large jaw crusher or a gyratory crusher, it can take nine months. _Source: Jim Redpath_ •The desired grizzly opening for an underground jaw crusher is equal to 80% of the gape of the crusher. _Source: Jack de la Vergne_ •The combination of a jaw crusher and a scalping grizzly will have 15% more capacity than a stand-alone jaw crusher. _Source: Ron Casson_ •As a rule, scalping grizzlies are rarely used anymore for (large) primary crushers. The exception is when ore contains wet fines that can cause acute packing in a gyratory crusher. _Source: McQuiston and Shoemaker_ •The product from a jaw crusher will tend to be less slabby and more even-dimensioned without a scalping grizzly, since slabs do not pass through so readily under this circumstance. _Source: A. L. Engels_ •Removal of the scalping grizzly for a primary jaw crusher can cut the liner life by 50%. It also makes it more difficult to clear a jam when the jaws are filled with fines. _Source: Ron Doyle_ 
*Area: *Rockbreakers •The capacity of a hydraulic rockbreaker is higher (and the operating cost lower) than a pneumatic rockbreaker. For these reasons, most new installations are hydraulic, despite the higher capital cost. _Source: John Kelly_ •For underground production rates less than 2,000 tpd, it may be economical to size the ore underground with rockbreakers only, otherwise, an underground crusher is usually necessary when skip hoisting is employed. _Source: John Gilbert_ •The operating cost for a stand-alone rockbreaker will be approximately 30% higher than it is for a crusher handling the same daily tonnage. _Source: John Gilbert_ •The capacity of one rockbreaker on a grizzly with the standard opening (± 16 by 18 inches) is in the order of 1,500-2,000 tpd. _Source: John Gilbert._ •For skips that fit into a standard 6 by 6 shaft compartment, the maximum particle size that is normally desired for skip hoisting is obtained when run-of-mine muck has been passed through a grizzly with a 16-18 inch opening. Skips hoisted in narrow shaft compartments may require a 12-14 inch spacing, while oversize skips may handle muck that has passed a 24-30 inch spacing. _Source: Jack de la Vergne_ •A pedestal mounted rockbreaker installed should be equipped with a boom that enables a reach of 20 feet (6m). _Source: Peter van Schaayk_


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## younes géol (30 أكتوبر 2011)

*thank you​*


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## alshangiti (6 نوفمبر 2011)

*Mining Field: *Ventilation and Air Conditioning 
*Area: *General •An underground trackless mine may require 10 tons of fresh air to be circulated for each ton of ore extracted. The hottest and deepest mines may use up to 20 tons of air for each ton of ore mined. _Source: Northern Miner Press_ •A factor of 100 cfm per ore-ton mined per day can be used to determine preliminary ventilation quantity requirements for most underground mining methods. Hot mines using ventilation air for cooling and mines with heavy diesel equipment usage require more air. Uranium mines require significantly higher ventilation quantities, up to 500 cfm per ton per day. Block cave and large-scale room and pillar mining operations require significantly lower ventilation quantities, in the range of 20 to 40 cfm per ton per day for preliminary calculations. _Source: Scott McIntosh_ •Ventilation is typically responsible for 40% of an underground mine's electrical power consumption. _Source: CANMET_ •If the exhaust airway is remote from the fresh air entry, approximately 85% of the fresh air will reach the intended destinations. If the exhaust airway is near to the fresh air entry, this can be reduced to 75%, or less. The losses are mainly due to leaks in ducts, bulkheads, and ventilation doors. _Source: Jack de la Vergne_ •Mine Resistance - For purposes of preliminary calculations, the resistance across the mine workings between main airway terminals underground (shafts, raises, air drifts, etc.) may be taken equal to one-inch water gauge. _Source: Richard Masuda_ •Natural pressure may be estimated at 0.03 inches of water gage per 10 degrees Fahrenheit difference per 100 feet difference in elevation (at standard air density). _Source: Robert Peele_ 
*Area: *Airways •The maximum practical velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,500 fpm (12.7m/s). _Source: Richard Masuda_ •The economic velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,400 fpm (12m/s). If the shaft incorporates a man-way compartment (ladder way) the economic velocity is reduced to about 1,400 fpm (7m/s). _Source: A.W.T. Barenbrug_ •The maximum velocity that should be contemplated for ventilation air in a circular concrete production shaft equipped with rope guides is 2,000 fpm and the recommended maximum relative velocity between skips and airflow is 6,000 fpm. _Source: Malcom McPherson_ •The "not-to-exceed" velocity for ventilation air in a bald circular concrete ventilation shaft is 4,000 fpm (20m/s). _Source: Malcom McPherson_ •The typical velocity for ventilation air in a bald circular concrete ventilation shaft or a bored raise is in the order of 3,200 fpm (16m/s) to be economical and the friction factor, k, is normally between 20 and 25. _Source: Jack de la Vergne_ •The typical velocity for ventilation air in a large raw (unlined) ventilation raise or shaft is in the order of 2,200 fpm (11m/s) to be economical and the friction factor, k, is typically between 60 and 75. _Source: Jack de la Vergne_ •The typical range of ventilation air velocities found in a conveyor decline or drift is between 500 and 1,000 fpm. It is higher if the flow is in the direction of conveyor travel and is lower against it. _Source: Floyd Bossard_ •The maximum velocity at draw points and dumps is 1,200 fpm (6m/s) to avoid dust entrainment. _Source: John Shilabeer_ •A protuberance into a smooth airway will typically provide four to five times the resistance to airflow as will an indent of the same dimensions. _Source: van den Bosch and Drummond_ •The friction factor, k, is theoretically constant for the same roughness of wall in an airway, regardless of its size. In fact, the factor is slightly decreased when the cross-section is large. _Source: George Stewart_ 
*Area: *Ducts •For bag duct, limiting static pressure to approximately 8 inches water gage will restrict leakage to a reasonable level. _Source: Bart Gilbert_ •The head loss of ventilation air flowing around a corner in a duct is reduced to 10% of the velocity head with good design. For bends up to 30 degrees, a standard circular arc elbow is satisfactory. For bends over 30 degrees, the radius of curvature of the elbow should be three times the diameter of the duct unless turning vanes inside the duct are employed. _Source: H.S. Fowler_ •The flow of ventilation air in a duct that is contracted will remain stable because the air-flow velocity is accelerating. The flow of ventilation air in a duct that is enlarged in size will be unstable unless the expansion is abrupt (high head loss) or it is coned at an angle of not more than 10 degrees (low head loss). _Source: H. S. Fowler_ 
*Area: *Fans •Increasing fan speed by 10% may increase the quantity of air by 10%, but the power requirement will increase by 33%. _Source: Chris Hall_ •The proper design of an evasée (fan outlet) requires that the angle of divergence not exceed 7 degrees. _Source: William Kennedy_ 
*Area: *Air Surveys •For a barometric survey, the correction factor for altitude may be assumed to be 1.11 kPa/100m (13.6 inches water gage per thousand feet). _Source: J.H. Quilliam_ 
*Area: *Clearing Smoke •The fumes from blasting operations cannot be removed from a stope or heading at a ventilation velocity less than 25 fpm (0.13m/s). A 30% higher air velocity is normally required to clear a stope. At least a 100% higher velocity is required to efficiently clear a long heading. _Source: William Meakin_ •The outlet of a ventilation duct in a development heading should be advanced to within 20 duct diameters of the face to ensure it is properly swept with fresh air. _Source: J.P. Vergunst_ •For sinking shallow shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 50 fpm (0.25m/s). _Source: Richard Masuda_ •For sinking deep shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 100 fpm (0.50m/s). _Source: Jack de la Vergne_ •For sinking very deep shafts, it is usually not practical to wait for smoke to clear. Normally, the first bucket of men returning to the bottom is lowered (rapidly) through the smoke. _Source: Morris Medd_ 
*Area: *Mine Air Heating •To avoid icing during winter months, a downcast hoisting shaft should have the air heated to at least 50C. (410 F.). A fresh air raise needs only 1.50C. (350 F.). _Source: Julian Kresowaty_ •When calculating the efficiency of heat transfer in a mine air heater, the following efficiencies may be assumed.
90% for a direct fired heater using propane, natural gas or electricity
80% for indirect heat transfer using fuel oil
_Source: Various_ •When the mine air is heated directly, it is important to maintain a minimum air stream velocity of approximately 2,400 fpm across the burners for efficient heat transfer. If the burners are equipped with combustion fans, lower air speeds (1,000 fpm) can be used. _Source: Andy Pitz_ •When the mine air is heated electrically, it is important to maintain a minimum air stream velocity of 400 fpm across the heaters. Otherwise, the elements will overheat and can burn out. _Source: Ed Summers_ 
*Area: *Heat Load •The lowest accident rates have been related to men working at temperatures below 70 degrees F and the highest to temperatures of 80 degrees and over. _Source: MSHA_ •Auto compression raises the dry bulb temperature of air by about 1 degree Celsius for every 100m the air travels down a dry shaft. (Less in a wet shaft.) The wet bulb temperature rises by approximately half this amount. _Source: Various_ •At depths greater than 2,000m, the heat load (due to auto compression) in the incoming air presents a severe problem. At these depths, refrigeration is required to remove the heat load in the fresh air as well as to remove the geothermal heat pick-up. _Source: Noel Joughin_ •At a rock temperature of 50 degrees Celsius, the heat load into a room and pillar stope is about 2.5 kW per square meter of face. _Source: Noel Joughin_ •In a hot mine, the heat generated by the wall rocks of permanent airways decays exponentially with time – after several months it is nearly zero. There remains some heat generated in permanent horizontal airways due to friction between the air and the walls. _Source: Jack de la Vergne_ •A diesel engine produces 200 cubic feet of exhaust gases per Lb. of fuel burned and consumption is approximately 0.45 Lb. of fuel per horsepower-hour. _Source: Caterpillar and others_ •Normally, the diesel engine on an LHD unit does not run at full load capacity (horsepower rating); it is more in the region of 50%, on average. In practice, all the power produced by the diesel engines of a mobile equipment fleet is converted into heat and each horsepower utilized produces heat equivalent to 42.4 BTU per minute. _Source: A.W.T.Barenbrug_ •The heat load from an underground truck or LHD is approximately 2.6 times as much for a diesel engine drive as it is for electric. _Source: John Marks_ •The efficiency of a diesel engine can be as high as 40% at rated RPM and full load, while that of an electric motor to replace it is as high as 96% at full load capacity. In both cases, the efficiency is reduced when operating at less than full load. _Source: Various_ •Normally, the electric motor on an underground ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, all the power produced by the electric motor of a booster fan or development heading fan is converted into heat and each horsepower (33,000 foot-Lb./minute) produces heat equivalent to 42.4 BTU per minute. (1 BTU = 778 foot-Lbs.) _Source: Jack de la Vergne_ •Normally, the electric motor on a surface ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, about 60% of the power produced by the electric motors of all the surface ventilation fans (intake and exhaust) is used to overcome friction in the intake airways and mine workings (final exhaust airways are not considered). Each horsepower lost to friction (i.e. static head) is converted into heat underground. _Source: Jack de la Vergne_ •Heat generated by electrically powered machinery underground is equal to the total power minus the motive power absorbed in useful work. The only energy consumed by electric motors that does not result in heat is that expended in work against gravity, such as hoisting, conveying up grade, or pumping to a higher elevation. _Source: Laird and Harris_ 
*Area: *Air Conditioning and Refrigeration •In the Republic of South Africa, cooling is required when the natural rock temperature reaches the temperature of the human body (98.6 degrees F). _Source: A.W.T. Barenbrug_ •A rough approximation of the cooling capacity required for a hot mine in North America is that the TR required per ton mined per day is 0.025 times the difference between the natural rock temperature (VRT) and 95 degrees F. For example, a 2,000 ton per day mine with a VRT of 140 degrees F. at the mean mining depth will require approximately 0.025 x 45 x 2,000 = 2,250 TR. _Source: Jack de la Vergne_ •The cold well (surge tank) for chilled surface water should have a capacity equal to the consumption of one shift underground. _Source: J. van der Walt_ •At the Homestake mine, the cost of mechanical refrigeration was approximately equal to the cost of ventilation. _Source: John Marks_


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## alshangiti (13 ديسمبر 2011)

Mineral Economics
*Area: *Metal Price•The long-term average price of a common mineral commodity (the price best used for economic evaluation in a feasibility study) is 1.5 times the average cost of production, worldwide. _Source: Sir Ronald Prain_
*Area: *Pre-production Capital Cost•The pre-production capital cost estimate (Capex) should include all construction and operating expenses until the mine has reached full production capacity or three months after reaching 50% of full capacity, whichever occurs first. This is the basic transition point between capital and operating costs. _Source: John Halls_•The pre-production capital cost expenditure includes all costs of construction and mine development until three months after the mine has reached 25% of its rated production capacity. _Source: Jon Gill_
*Area: *Cash Flow•The total cash flow must be sufficient to repay the capital cost at least twice. _Source: L. D. Smith_•Project loans should be repaid before half the known reserves are consumed. _Source: G.R Castle_•Incremented cash flow projections should each be at least 150% of the loan repayment scheduled for the same period. _Source: G.R. Castle_•The operating cost should not exceed half the market value of minerals recovered. _Source: Alan Provost_
*Area: *Net Present Value•The discount factor employed to determine the NPV is often 10%; however, it should be Prime + 5%. _Source: G.R. Castle_•In numerous conversations with managers of mining firms, I have found that 15% in real terms is the common discount rate used for decision purposes. _Source: Herbert Drecshler (1980)_•The true present value (market value) of a project determined for purposes of joint venture or outright purchase is equal to half the NPV typically calculated. _Source: J. B. Redpath_
*Area: *Rate of Return•The feasibility study for a hard rock mine should demonstrate an internal rate of return (IRR) of at least 20% – more during periods of high inflation. _Source: J. B. Redpath_
*Area: *Working Capital•Working capital equals ten weeks operating cost plus cost of capital spares and parts. _Source: Alan O'Hara_
*Area: *Closure Costs•The salvage value of plant and equipment should pay for the mine closure costs. _Source: Ron Haflidson_


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## alshangiti (7 يونيو 2012)

*	•	An allowance (such as 15%) should be specifically determined and added to the contractor's formal bid price for a mining project to account for contract clauses relating to unavoidable extra work, delays, ground conditions, over-break, grouting, de-watering, claims, and other unforeseen items. Source: Jack de la Vergne
*	
Area: Engineering, Procurement, and Construction Management
*	•	The Engineering, Procurement, and Construction Management (EPCM) cost will be approximately 17% for surface and underground construction and 5% for underground development. Source: Jack de la Vergne
*	
Area: Overbreak
*	•	The amount of over-break to be estimated against rock for a concrete pour will average approximately one foot in every applicable direction, more at brows, lips, and in bad ground. Source: Jack de la Vergne
*	•	On average, for each one cubic yard of concrete measured from the neat lines on drawings, there will be two cubic yards required underground, due to over-break and waste. Source: Jack de la Vergne
*	
Area: Haulage
*	•	The economical tramming distance for a 5 cubic yard capacity LHD is 500 feet and will produce 500 tons per shift, for an 8-yard LHD, it is 800 feet and 800 tons per shift. Source: Sandy Watson
*	•	Haulage costs for open pit are at least 40% of the total mining costs; therefore, proximity of the waste dumps to the rim of the pit is of great importance. Source: Frank Kaeschager
*	
Area: Miscellaneous
*	•	The installed cost of a long conveyorway is approximately equal to the cost of driving the drift or decline in which it is to be placed. Source: Jack de la Vergne
*	•	In a trackless mine operating around the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. Source: John Gilbert
*	•	On average, for each cubic yard of concrete measured from the neat lines on drawings, approximately 110 Lbs. of reinforcing steel and 12 square feet of forms will be required. Source: Jack de la Vergne
*	•	The overall advance rate of a trackless heading may be increased by 30% and the unit cost decreased by 15% when two headings become available. Source: Bruce Lang
*	•	The cost to slash a trackless heading wider while it is being advanced is 80% of the cost of the heading itself, on a volumetric basis. Source: Bruce Lang


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## alshangiti (7 يونيو 2012)

Mining Field: Cost Estimating
*	
Area: Cost of Estimating
*	•	A detailed estimate for routine, repetitive work (i.e. a long drive on a mine level) may cost as little as 0.5% of the project cost. On the other hand, it may cost up to 5% to adequately estimate projects involving specialized work, such as underground construction and equipment installation. Source: Various
*	
Area: Cost of Feasibility Study
*	•	The cost of a detailed feasibility study will be in a range from 0.5% to 1.5% of the total estimated project cost. Source: Frohling and Lewis
*	•	The cost of a detailed or "bankable" feasibility study is typically in the range of 2% to 5% of the project, if the costs of additional (in-fill) drilling, assaying, metallurgical testing, geotechnical investigations, etc. are added to the direct and indirect costs of the study itself. Source: R. S. Frew
*	
Area: Budget Estimates
*	•	An allowance (such as 15%) should be specifically determined and added to the contractor's formal bid price for a mining project to account for contract clauses relating to unavoidable extra work, delays, ground conditions, over-break, grouting, de-watering, claims, and other unforeseen items. Source: Jack de la Vergne


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## alshangiti (7 يونيو 2012)

Back fill. 


Mining Field: Backfill
*	
Area: General
*	•	The cost of backfilling will be near 20% of the total underground operating cost. Source: Bob Rappolt
*	•	The capital cost of a paste fill plant installation is approximately twice the cost of a conventional hydraulic fill plant of the same capacity. Source: Barrett, Fuller, and Miller
*	•	If a mine backfills all production stopes to avoid significant delays in ore production, the daily capacity of the backfill system should be should be at least 1.25 times the average daily mining rate (expressed in terms of volume). Source: Robert Currie
*	•	The typical requirement for backfill is approximately 50% of the tonnage mined. It is theoretically about 60%, but all stopes are not completely filled and tertiary stopes may not be filled at all. Source: Ross Gowan
*	•	It is common to measure the strength of cemented backfill as if it were concrete (i.e. 28 days), probably because this time coincides with the planned stope turn-around cycle. Here it should be noted that while concrete obtains over 80% of its long- term strength at 28 days, cemented fill might only obtain 50%. In other words, a structural fill may have almost twice the strength at 90 days as it had at 28 days. Source: Jack de la Vergne
*	
Area: Hydraulic Fill
*	•	Because the density of hydraulic fill when placed is only about half that of ore, unless half the tailings can be recovered to meet gradation requirements, a supplementary or substitute source of fill material is required. Source: E. G. Thomas
*	
Area: Cemented Rock Fill
*	•	A 6% binder will give almost the same CRF strength in 14 days that a 5% binder will give in 28 days. This rule is useful to know when a faster stope turn-around time becomes necessary. Source: Joel Rheault
*	•	As the fly ash content of a CRF slurry is increased above 50%, the strength of the backfill drops rapidly and the curing time increases dramatically. A binder consisting of 35% fly ash and 65% cement is deemed to be the optimal mix. Source: Joel Rheault
*	•	The size of water flush for a CRF slurry line should be 4,000 US gallons. Source: George Greer
*	•	The optimum W/C ratio for a CRF slurry is 0.8:1, but in practice, the water content may have to be reduced when the rock is wet due to ice and snow content of quarried rock or ground water seepage into the fill raise. Source: Finland Tech
*	•	The actual strength of CRF placed in a mine will be approximately 2/3 the laboratory value that is obtained from standard 6 inch diameter concrete test cylinders, but will be about 90% of the value obtained from 12 inch diameter cylinders. Source: Thiann Yu
*	
Area: Paste Fill
*	•	Only about 60% of mill tailings can be used for paste fill over the life of a mine because of the volume increase, which occurs as a result of breaking and comminuting the ore. Source: David Landriault
*	•	Experience to date at the Golden Giant mine indicates that only 46% of the tailings produced can be used for paste fill. Source: Jim Paynter
*	•	Very precise control of pulp density is required for gravity flow of paste fill. A small (1-2%) increase in pulp density can more than double pipeline pressures (and resistance to flow). Source: David Landriault


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## alshangiti (20 يونيو 2012)

Sinking shaft. 

*	
Area: Schedule
*	•	From time of award to the start of sinking a timber shaft will be approximately five months. A circular concrete shaft may take three months longer unless the shaft collar and headframe are completed in advance. Source: Tom Anderson
*	•	The average rate of advance for shaft sinking will be two-thirds of the advance in the best month (the one everyone talks about). Source: Jim Redpath
*	
Area: Hoist
*	•	The hoist required for shaft sinking needs approximately 30% more horsepower than for skipping the same payload at the same line speed. Source: Jack de la Vergne
*	•	Without slowing the rate of advance, a single drum hoist is satisfactory to sink to a depth of 1,500 feet at five buckets per foot, 2,000 feet at four buckets per foot, and 2,500 feet at 3½ buckets per foot. For deeper shafts, a double drum hoist is required to keep up with the shaft mucker. Source: Jack de la Vergne
*	
Area: Bucket
*	•	For sinking a vertical shaft, the bucket size should be at least big enough to fill six for each foot of shaft to be sunk; five is better. Source: Marshall Hamilton
*	•	For the bucket to remain stable when detached on the shaft bottom, its height should not exceed its diameter by more than 50%. Source: Jim Redpath
*	•	Tall buckets can be used safely if the clam is used to dig a hole in the muck pile for the buckets. Source: Bill Shaver
*	•	A bucket should not be higher than 7½ feet for filling with a standard Cryderman clam (which has an 11-foot stroke). Source: Bert Trenfield
*	•	A bucket should not be higher than 6 feet when mucking with a 630, which has a 6-foot-6-inch discharge height. Source: Alan Provost
*	•	You can load a tall bucket using a 630 if you slope the muck pile so that the bucket sits at an angle from the vertical position. Source: Fern Larose
*	•	In a wet shaft, the contractor should be able to bail up to 10 buckets of water per shift without impeding his advance. Source: Paddy Harrison
*	
Area: Water Pressure
*	•	For any shaft, the water pressure reducing valves should be installed every 250 feet. "Toilet tank" reducers are more reliable than valves and may be spread further apart. Source: Peter van Schaayk
*	•	Water pressure reducing valves may be eliminated for shaft sinking if the water line is slotted and the drill water is fed in batch quantities. Source: Allan Widlake and Jannie Mostert
*	
Area: Compressed Air
*	•	One thousand cfm of compressed air is needed to blow the bench with a two-inch blowpipe. Source: Bill Shaver
*	•	Twelve hundred cfm of compressed air is needed to operate a standard Cryderman clam properly. Source: Bill Shaver
*	
Area: Shaft Stations
*	•	The minimum station depth at a development level to be cut during shaft sinking is 50 feet. Source: Tom Goodell
*	•	A shaft station will not be cut faster than 2,000 cubic feet per day with slusher mucking. It may be cut at an average rate of 3,500 cubic feet per day with an LHD mucking unit. Source: Jim Redpath
*	
Area: Circular Shaft
*	•	The minimum (finished) diameter of a circular shaft for bottom mucking with a 630-crawler loader is 18 feet. Source: Tom Goodell
*	•	For a circular concrete shaft, the minimum clearance between the sinking stage and the shaft walls is 10 inches. Source: Henry Lavigne
*	•	A circular concrete lined shaft sunk in good ground will have an average overbreak of 10 inches or more, irrespective of the minimum concrete thickness. Source: Jim Redpath
*	•	For a rope guide system in a shaft being sunk to a moderate depth, the minimum clearance between a conveyance (bucket and crosshead) and a fixed obstruction is 12 inches and to another bucket is 24 inches. At the shaft collar, the clearance to a fixed obstruction may be reduced to 6 inches due to slowdown, or less with the use of fairleads or skid plates. In a deep shaft, 18-24 inches is required to clear a fixed obstruction and 30-36 inches is required between buckets, depending on the actual hoisting speed. These clearances assume that the shaft stage hangs free and the guide ropes are fully tensioned when hoisting buckets. Source: Various
*	•	When hoisting at speeds approaching 3,000 fpm (15m/s) on a rope guide system, the bonnet of the crosshead should be grilled instead of being constructed of steel plate to minimize aerodynamic sway. Source: Morris Medd
*	•	The maximum rate at which ready-mix concrete will be poured down a 6-inch diameter slick line is 60 cubic yards per hour. Source: Marshall Hamilton
*	
Area: Timber Shaft
*	•	For a timber shaft, the minimum clearance to the wall rock outside wall plates and end plates should be 6 inches; the average will be 14 inches in good ground. Source: Alan Provost
*	•	For a timber shaft that encounters squeezing ground, the minimum clearance outside wall plates and end plates should be 12 inches. Source: Dan Hinich
*	•	For a timber shaft, the blocking should not be longer than two feet without being pinned with rock bolts to the wall rock. Source: Jim Redpath


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## alshangiti (16 يوليو 2012)

*Mining Field: *Backfill  
*Area: *General  •The cost of backfilling will be near 20% of the total underground operating cost. _Source: Bob Rappolt_ •The capital cost of a paste fill plant installation is approximately twice the cost of a conventional hydraulic fill plant of the same capacity. _Source: Barrett, Fuller, and Miller_ •If a mine backfills all production stopes to avoid significant delays in ore production, the daily capacity of the backfill system should be should be at least 1.25 times the average daily mining rate (expressed in terms of volume). _Source: Robert Currie_ •The typical requirement for backfill is approximately 50% of the tonnage mined. It is theoretically about 60%, but all stopes are not completely filled and tertiary stopes may not be filled at all. _Source: Ross Gowan_ •It is common to measure the strength of cemented backfill as if it were concrete (i.e. 28 days), probably because this time coincides with the planned stope turn-around cycle. Here it should be noted that while concrete obtains over 80% of its long- term strength at 28 days, cemented fill might only obtain 50%. In other words, a structural fill may have almost twice the strength at 90 days as it had at 28 days. _Source: Jack de la Vergne_ 
*Area: *Hydraulic Fill  •Because the density of hydraulic fill when placed is only about half that of ore, unless half the tailings can be recovered to meet gradation requirements, a supplementary or substitute source of fill material is required. _Source: E. G. Thomas_ 
*Area: *Cemented Rock Fill  •A 6% binder will give almost the same CRF strength in 14 days that a 5% binder will give in 28 days. This rule is useful to know when a faster stope turn-around time becomes necessary. _Source: Joel Rheault_ •As the fly ash content of a CRF slurry is increased above 50%, the strength of the backfill drops rapidly and the curing time increases dramatically. A binder consisting of 35% fly ash and 65% cement is deemed to be the optimal mix. _Source: Joel Rheault_ •The size of water flush for a CRF slurry line should be 4,000 US gallons. _Source: George Greer_ •The optimum W/C ratio for a CRF slurry is 0.8:1, but in practice, the water content may have to be reduced when the rock is wet due to ice and snow content of quarried rock or ground water seepage into the fill raise. _Source: Finland Tech_ •The actual strength of CRF placed in a mine will be approximately 2/3 the laboratory value that is obtained from standard 6 inch diameter concrete test cylinders, but will be about 90% of the value obtained from 12 inch diameter cylinders. _Source: Thiann Yu_ 
*Area: *Paste Fill  •Only about 60% of mill tailings can be used for paste fill over the life of a mine because of the volume increase, which occurs as a result of breaking and comminuting the ore. _Source: David Landriault_ •Experience to date at the Golden Giant mine indicates that only 46% of the tailings produced can be used for paste fill. _Source: Jim Paynter_ •Very precise control of pulp density is required for gravity flow of paste fill. A small (1-2%) increase in pulp density can more than double pipeline pressures (and resistance to flow). _Source: David Landriault_ •40% of paste fill distribution piping may be salvaged for re-use. _Source: BM&S Corporation_


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## alshangiti (13 سبتمبر 2012)

*
Mining Field: Exploration Geology and Ore Reserves
*	
Area: Discovery
*	•	It takes 25,000 claims staked to find 500 worth diamond drilling to find one mine. Source: Lorne Ames
*	•	On average, the time between discovery and actual start of construction of a base metal mine is 10 years; it is less for a precious metal mine. Source: J.P. Albers
*	
Area: Costs
*	•	The amount expended on diamond drilling and exploration development for the purposes of measuring mineral reserves should approximately equal 2% of the gross value of the metals in the reserves. Source: Joe Gerden
*	
Area: Bulk Sample
*	•	The minimum size of a bulk sample, when required for a proposed major open pit mine, is in the order of 50,000 tons (with a pilot mill on site). For a proposed underground mine, it is typically only 5,000 tons. Source: Jack de la Vergne
*	
Area: Ore Reserve Estimate
*	•	The value reported for the specific gravity (SG) of an ore sample on a metallurgical test report is approximately 20% higher than the correct value to be employed in the resource tonnage calculation. Source: Jack de la Vergne
*	
Area: Ore Resource Estimate
*	•	To determine "inferred" or "possible" reserves, it is practice to assume that the ore will extend to a distance at least equal to half the strike length at the bottom of measured reserves. Another rule is that the largest horizontal cross section of an ore body is half way between its top and bottom. Source: H. E. McKinstry
*	•	In the base metal mines of Peru and the Canadian Shield, often a zonal mineralogy is found indicating depth. At the top of the ore body sphalerite and galena predominate. Near mid-depth, chalcopyrite becomes significant and pyrite appears. At the bottom, pyrite, and magnetite displace the ore. Source: H. E. McKinstry
*	•	In gold mines, the amount of silver that accompanies the gold may be an indicator of depth. Shallow gold deposits usually have relatively high silver content while those that run deep have hardly any. Source: James B. Redpath
*	•	As a rule of thumb, I use that 2P (Probable) reserves are only such when drill spacing does not exceed five to seven smallest mining units (SMU). Open pit mining on 15m benches could have an SMU of 15m by 15m by 15m. Underground, an SMU would be say 3m by 3m by 3m (a drift round). Source: René Marion
*	
Area: Strike and Dip
*	•	The convention for establishing strike and dip is always the Right Hand Rule. With right hand palm up, open and extended, point the thumb in the down-dip direction and the fingertips provide the strike direction. Source: Mike Neumann


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## alshangiti (28 سبتمبر 2012)

*	
Area: General
*	•	An underground trackless mine may require 10 tons of fresh air to be circulated for each ton of ore extracted. The hottest and deepest mines may use up to 20 tons of air for each ton of ore mined. Source: Northern Miner Press
*	•	A factor of 100 cfm per ore-ton mined per day can be used to determine preliminary ventilation quantity requirements for most underground mining methods. Hot mines using ventilation air for cooling and mines with heavy diesel equipment usage require more air. Uranium mines require significantly higher ventilation quantities, up to 500 cfm per ton per day. Block cave and large-scale room and pillar mining operations require significantly lower ventilation quantities, in the range of 20 to 40 cfm per ton per day for preliminary calculations. Source: Scott McIntosh
*	•	Ventilation is typically responsible for 40% of an underground mine's electrical power consumption. Source: CANMET
*	•	If the exhaust airway is remote from the fresh air entry, approximately 85% of the fresh air will reach the intended destinations. If the exhaust airway is near to the fresh air entry, this can be reduced to 75%, or less. The losses are mainly due to leaks in ducts, bulkheads, and ventilation doors. Source: Jack de la Vergne
*	•	Mine Resistance - For purposes of preliminary calculations, the resistance across the mine workings between main airway terminals underground (shafts, raises, air drifts, etc.) may be taken equal to one-inch water gauge. Source: Richard Masuda
*	•	Natural pressure may be estimated at 0.03 inches of water gage per 10 degrees Fahrenheit difference per 100 feet difference in elevation (at standard air density). Source: Robert Peele
*	
Area: Airways
*	•	The maximum practical velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,500 fpm (12.7m/s). Source: Richard Masuda
*	•	The economic velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,400 fpm (12m/s). If the shaft incorporates a man-way compartment (ladder way) the economic velocity is reduced to about 1,400 fpm (7m/s). Source: A.W.T. Barenbrug
*	•	The maximum velocity that should be contemplated for ventilation air in a circular concrete production shaft equipped with rope guides is 2,000 fpm and the recommended maximum relative velocity between skips and airflow is 6,000 fpm. Source: Malcom McPherson
*	•	The "not-to-exceed" velocity for ventilation air in a bald circular concrete ventilation shaft is 4,000 fpm (20m/s). Source: Malcom McPherson
*	•	The typical velocity for ventilation air in a bald circular concrete ventilation shaft or a bored raise is in the order of 3,200 fpm (16m/s) to be economical and the friction factor, k, is normally between 20 and 25. Source: Jack de la Vergne
*	•	The typical velocity for ventilation air in a large raw (unlined) ventilation raise or shaft is in the order of 2,200 fpm (11m/s) to be economical and the friction factor, k, is typically between 60 and 75. Source: Jack de la Vergne
*	•	The typical range of ventilation air velocities found in a conveyor decline or drift is between 500 and 1,000 fpm. It is higher if the flow is in the direction of conveyor travel and is lower against it. Source: Floyd Bossard
*	•	The maximum velocity at draw points and dumps is 1,200 fpm (6m/s) to avoid dust entrainment. Source: John Shilabeer
*	•	A protuberance into a smooth airway will typically provide four to five times the resistance to airflow as will an indent of the same dimensions. Source: van den Bosch and Drummond
*	•	The friction factor, k, is theoretically constant for the same roughness of wall in an airway, regardless of its size. In fact, the factor is slightly decreased when the cross-section is large. Source: George Stewart
*	
Area: Ducts
*	•	For bag duct, limiting static pressure to approximately 8 inches water gage will restrict leakage to a reasonable level. Source: Bart Gilbert
*	•	The head loss of ventilation air flowing around a corner in a duct is reduced to 10% of the velocity head with good design. For bends up to 30 degrees, a standard circular arc elbow is satisfactory. For bends over 30 degrees, the radius of curvature of the elbow should be three times the diameter of the duct unless turning vanes inside the duct are employed. Source: H.S. Fowler
*	•	The flow of ventilation air in a duct that is contracted will remain stable because the air-flow velocity is accelerating. The flow of ventilation air in a duct that is enlarged in size will be unstable unless the expansion is abrupt (high head loss) or it is coned at an angle of not more than 10 degrees (low head loss). Source: H. S. Fowler
*	
Area: Fans
*	•	Increasing fan speed by 10% may increase the quantity of air by 10%, but the power requirement will increase by 33%. Source: Chris Hall
*	•	The proper design of an evasée (fan outlet) requires that the angle of divergence not exceed 7 degrees. Source: William Kennedy
*	
Area: Air Surveys
*	•	For a barometric survey, the correction factor for altitude may be assumed to be 1.11 kPa/100m (13.6 inches water gage per thousand feet). Source: J.H. Quilliam
*	
Area: Clearing Smoke
*	•	The fumes from blasting operations cannot be removed from a stope or heading at a ventilation velocity less than 25 fpm (0.13m/s). A 30% higher air velocity is normally required to clear a stope. At least a 100% higher velocity is required to efficiently clear a long heading. Source: William Meakin
*	•	The outlet of a ventilation duct in a development heading should be advanced to within 20 duct diameters of the face to ensure it is properly swept with fresh air. Source: J.P. Vergunst
*	•	For sinking shallow shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 50 fpm (0.25m/s). Source: Richard Masuda
*	•	For sinking deep shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 100 fpm (0.50m/s). Source: Jack de la Vergne
*	•	For sinking very deep shafts, it is usually not practical to wait for smoke to clear. Normally, the first bucket of men returning to the bottom is lowered (rapidly) through the smoke. Source: Morris Medd
*	
Area: Mine Air Heating
*	•	To avoid icing during winter months, a downcast hoisting shaft should have the air heated to at least 50C. (410 F.). A fresh air raise needs only 1.50C. (350 F.). Source: Julian Kresowaty
*	•	When calculating the efficiency of heat transfer in a mine air heater, the following efficiencies may be assumed.
90% for a direct fired heater using propane, natural gas or electricity
80% for indirect heat transfer using fuel oil
Source: Various
*	•	When the mine air is heated directly, it is important to maintain a minimum air stream velocity of approximately 2,400 fpm across the burners for efficient heat transfer. If the burners are equipped with combustion fans, lower air speeds (1,000 fpm) can be used. Source: Andy Pitz
*	•	When the mine air is heated electrically, it is important to maintain a minimum air stream velocity of 400 fpm across the heaters. Otherwise, the elements will overheat and can burn out. Source: Ed Summers
*	
Area: Heat Load
*	•	The lowest accident rates have been related to men working at temperatures below 70 degrees F and the highest to temperatures of 80 degrees and over. Source: MSHA
*	•	Auto compression raises the dry bulb temperature of air by about 1 degree Celsius for every 100m the air travels down a dry shaft. (Less in a wet shaft.) The wet bulb temperature rises by approximately half this amount. Source: Various
*	•	At depths greater than 2,000m, the heat load (due to auto compression) in the incoming air presents a severe problem. At these depths, refrigeration is required to remove the heat load in the fresh air as well as to remove the geothermal heat pick-up. Source: Noel Joughin
*	•	At a rock temperature of 50 degrees Celsius, the heat load into a room and pillar stope is about 2.5 kW per square meter of face. Source: Noel Joughin
*	•	In a hot mine, the heat generated by the wall rocks of permanent airways decays exponentially with time � after several months it is nearly zero. There remains some heat generated in permanent horizontal airways due to friction between the air and the walls. Source: Jack de la Vergne
*	•	A diesel engine produces 200 cubic feet of exhaust gases per Lb. of fuel burned and consumption is approximately 0.45 Lb. of fuel per horsepower-hour. Source: Caterpillar and others
*	•	Normally, the diesel engine on an LHD unit does not run at full load capacity (horsepower rating); it is more in the region of 50%, on average. In practice, all the power produced by the diesel engines of a mobile equipment fleet is converted into heat and each horsepower utilized produces heat equivalent to 42.4 BTU per minute. Source: A.W.T.Barenbrug
*	•	The heat load from an underground truck or LHD is approximately 2.6 times as much for a diesel engine drive as it is for electric. Source: John Marks
*	•	The efficiency of a diesel engine can be as high as 40% at rated RPM and full load, while that of an electric motor to replace it is as high as 96% at full load capacity. In both cases, the efficiency is reduced when operating at less than full load. Source: Various
*	•	Normally, the electric motor on an underground ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, all the power produced by the electric motor of a booster fan or development heading fan is converted into heat and each horsepower (33,000 foot-Lb./minute) produces heat equivalent to 42.4 BTU per minute. (1 BTU = 778 foot-Lbs.) Source: Jack de la Vergne
*	•	Normally, the electric motor on a surface ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, about 60% of the power produced by the electric motors of all the surface ventilation fans (intake and exhaust) is used to overcome friction in the intake airways and mine workings (final exhaust airways are not considered). Each horsepower lost to friction (i.e. static head) is converted into heat underground. Source: Jack de la Vergne
*	•	Heat generated by electrically powered machinery underground is equal to the total power minus the motive power absorbed in useful work. The only energy consumed by electric motors that does not result in heat is that expended in work against gravity, such as hoisting, conveying up grade, or pumping to a higher elevation. Source: Laird and Harris
*	
Area: Air Conditioning and Refrigeration
*	•	In the Republic of South Africa, cooling is required when the natural rock temperature reaches the temperature of the human body (98.6 degrees F). Source: A.W.T. Barenbrug
*	•	A rough approximation of the cooling capacity required for a hot mine in North America is that the TR required per ton mined per day is 0.025 times the difference between the natural rock temperature (VRT) and 95 degrees F. For example, a 2,000 ton per day mine with a VRT of 140 degrees F. at the mean mining depth will require approximately 0.025 x 45 x 2,000 = 2,250 TR. Source: Jack de la Vergne


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## alshangiti (23 أكتوبر 2012)

Lateral Development and Ramps  
*Area: *General  •Laser controls should be used in straight development headings that exceed 800 feet (240m) in length. _Source: Tom Goodell_ •The overall advance rate of a lateral drive may be increased by 30% and the unit cost decreased by 15% when two headings become available. _Source: Bruce Lang_ 
*Area: *Trackless Headings  •The minimum width for a trackless heading is 5 feet wider than the widest unit of mobile equipment. _Source: Fred Edwards_ •The back (roof) of trackless headings in hard rock should be driven with an arch of height equal to 20% of the heading width. _Source: Kidd Mine Standards_ •The cost to slash a trackless heading wider while it is being advanced is 80% of the cost of the heading itself, on a volumetric basis. _Source: Bruce Lang_ •For long ramp drives, the LHD/truck combination gives lower operating costs than LHDs alone and should be considered on any haul more than 1,500 feet in length. _Source: Jack Clark_ •LHD equipment is usually supplemented with underground trucks when the length of drive exceeds 1,000 feet. _Source: Fred Edwards_ •With ramp entry, a satellite shop is required underground for mobile drill jumbos and crawler mounted drills when the mean mining depth reaches 200m below surface. _Source: Jack de la Vergne_ •With ramp and shaft entry, a main shop is required underground when the mean mining depth reaches 500m below surface. _Source: Jack de la Vergne_ •A gradient of 2% is not enough for a horizontal trackless heading. It ought to be driven at a minimum of 2½% or 3%. _Source: Bill Shaver_ •The minimum radius of drift or ramp curve around which it is convenient to drive a mobile drill jumbo is 75 feet. _Source: Al Walsh_ •For practical purposes, a minimum curve radius of 50 feet may be employed satisfactorily for most ramp headings. _Source: John Gilbert_ •The gathering arm reach of a continuous face-mucking unit should be 2 feet wider than the nominal width of the drift being driven. _Source: Jim Dales_ •Footwall drifts for trackless blasthole mining should be offset from the ore by at least 15m (50 feet) in good ground. Deeper in the mine, the offset should be increased to 23m (75 feet) and for mining at great depth it should be not less than 30m (100 feet). _Source: Jack de la Vergne_ •Ore passes should be spaced at intervals not exceeding 500 feet (and waste passes not more than 750 feet) along the draw point drift, with LHD extraction. _Source: Jack de la Vergne_ •The maximum practical air velocity in lateral headings that are travelways is approximately 1,400 fpm. Even at this speed, a hard hat may be blown off when a vehicle or train passes by. At higher velocities, walking gets difficult and road dust becomes airborne. However, in pure lateral airways, the air velocity may exceed 3,000 fpm. _Source: Various_ •The typical range of ventilation air velocities found in a conveyor decline or drift is between 500 and 1,000 fpm. It is higher if the flow is in the direction of conveyor travel and is lower against it. _Source: Floyd Bossard_ •The maximum velocity at draw points and dumps is 1,200 fpm (6m/s) to avoid dust entrainment. _Source: John Shilabeer_ 
*Area: *Track Headings  •Track gage should not be less than ½ the extreme width of car or motor (locomotive). _Source: MAPAO_ •The tractive effort, TE (Lbs.) for a diesel locomotive is approximately equal to 300 times its horsepower rating. _Source: John Partridge_ •Wood ties should have a length equal to twice the track gage, be at least ¼ inch thicker than the spike length, and 1 3/8 times spike length in width. _Source: MAPAO_ •Typical gradients for track mines are 0.25% and 0.30%. _Source: MAPAO_ •A minimum clearance of three feet should be designed between the outside of the rails and the wall of the drift to permit safe operation of a mucking machine when driving the heading. _Source: MAPAO_


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## alshangiti (6 فبراير 2013)

*	
Area: Wood Headframe
*	•	The maximum height of a wood headframe is 110 feet. The maximum rope size for a wood headframe is 1.25 inches diameter, which corresponds to an 8-foot or 100-inch diameter double-drum hoist. Source: Jack de la Vergne
*	
Area: Steel Headframe
*	•	A headframe (for a ground mounted hoist) should be designed with the backlegs at an angle of 60 degrees from the horizontal and the rope flight from the hoist at an angle of 45 degrees. Source: Mine Plant Design, Staley, 1949
*	•	It is better to design a headframe (for a ground mounted hoist) such that the resultant of forces from the overwound rope falls about 1/3 the distance from the backleg to the backpost. Source: Mine Plant Design, Staley, 1949
*	•	No members in a steel headframe should have a thickness less than 5/16 of an inch. Main members should have a slenderness ratio (l/r) of not more than 120; secondary members not more than 200. Source: Mine Plant Design, Staley, 1949
*	•	Main members of a modern steel headframe may have a slenderness ratio as high as 160 meeting relevant design codes and modern design practice. Source: Steve Boyd
*	
Area: Steel Headframe versus Concrete Headframe
*	•	The cost of a steel headframe increases exponentially with its height while the cost of a concrete headframe is nearly a direct function of its height. As a result, a steel headframe is less expensive than a concrete headframe, when the height of the headframe is less than approximately 160 feet (at typical market costs for structural steel and ready-mix concrete). Source: Jack de la Vergne
*	•	At the hoist deck level of a tower mount headframe for Koepe hoisting, the maximum permissible lateral deflection (due to wind sway, foundation settlement, etc.) is 3 inches. (This may favor a concrete headframe.) Source: R. L. Puryear
*	•	A concrete headframe will weigh up to ten times as much as the equivalent steel headframe. (This may favor the steel headframe when foundations are in overburden or the mine site is in a seismic zone.) Source: Steve Boyd
*	
Area: Headframe Bins
*	•	To determine the live load of a surface bin for a hard rock mine, the angle of repose may be assumed at 35 degrees from the horizontal (top of bin) and the angle of drawdown assumed at 60 degrees. Source: Al Fernie
*	•	A bin for a hard rock mine will likely experience rat-holing (as opposed to mass flow) if the ore is damp, unless the dead bed at the bin bottom is covered or replaced with a smooth steel surface at an angle of approximately 60 degrees from the horizontal. Source: Jennike and Johanson
*	•	The live-load capacity of the headframe ore bin at a small mine (where trucking of the ore is employed) may be designed equal to a day's production. For a mine of medium size, it can be as little as one-third of a day's production. For a high capacity skipping operation, the headframe should have a conveyor load-out, either direct to the mill or elevated to separate load-out bins remote from the headframe. A conveyor load-out requires a small surge bin at the headframe of live load capacity approximately equal to the payload of 20 skips. Source: Various


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## alshangiti (16 مايو 2013)

Mining Field: Rock Mechanics

Area: Ground Stress
•	The vertical stress may be calculated on the basis of depth of overburden with an accuracy of ± 20%. This is sufficient for engineering purposes. Source: Z.T. Bieniawski
•	Discs occur in the core of diamond drill holes when the radial ground stresses are in excess of half the compressive rock strength. Source: Obert and Stephenson
•	The width of the zone of relaxed stress around a circular shaft that is sunk by a drill and blast method is approximately equal to one-third the radius of the shaft excavation. Source: J. F. Abel

Area: Ground Control
•	The length of a rock bolt should be one-half to one-third the heading width. Source: Mont Blanc Tunnel Rule (c.1965)
•	In hard rock mining, the ratio of bolt length to pattern spacing is normally 1½:1. In fractured rock, it should be at least 2:1. (In civil tunnels and coalmines, it is typically 2:1.) Source: Lang and Bischoff (1982)
•	In mining, the bolt length/bolt spacing ratio is acceptable between 1.2:1 and 1.5:1. Source: Z.T. Bieniawski (1992)
•	In good ground, the length of a roof bolt can be one-third of the span. The length of a wall bolt can be one-fifth of the wall height. The pattern spacing may be obtained by dividing the rock bolt length by one and one-half. Source: Mike Gray (1999)
•	The tension developed in a mechanical rock bolt is increased by approximately 40 Lbs. for each one foot-pound increment of torque applied to it. Source: Lewis and Clarke
•	A mechanical rock bolt installed at 30 degrees off the perpendicular may provide only 25% of the tension produced by a bolt equally torqued that is perpendicular to the rock face, unless a spherical washer is employed. Source: MAPAO
•	For each foot of friction bolt (split-set) installed, there is 1 ton of anchorage. Source: MAPAO
•	The shear strength (dowel strength) of a rock bolt may be assumed equal to one-half its tensile strength. Source: P. M. Dight
•	The thickness of the beam (zone of uniform compression) in the back of a bolted heading is approximately equal to the rock bolt length minus the spacing between them. Source: T.A. Lang
•	Holes drilled for resin bolts should be ¼ inch larger in diameter than the bolt. If it is increased to 3/8 inch, the pull out load is not affected but the stiffness of the bolt/resin assembly is lowered by more than 80%, besides wasting money on unnecessary resin. Source: Dr. Pierre Choquette
•	Holes drilled for cement-grouted bolts should be ½ to 1 inch larger in diameter than the bolt. The larger gap is especially desired in weak ground to increase the bonding area. Source: Dr. Pierre Choquette

Area: Mine Development
•	Permanent underground excavations should be designed to be in a state of compression. A minimum safety factor (SF) of 2 is generally recommended for them. Source: Obert and Duval
•	The required height of a rock pentice to be used for shaft deepening is equal to the shaft width or diameter plus an allowance of five feet. Source: Jim Redpath

Area: Stope Pillar and Design
•	A minimum SF of between 1.2 and 1.5 is typically employed for the design of rigid stope pillars in hard rock mines. Source: Various
•	For purposes of pillar design in hard rock, the uniaxial compressive strength obtained from core samples should be reduced by 20-25% to obtain a true value underground. The reduced value should be used when calculating pillar strength from formulas relating it to compressive strength, pillar height, and width (i.e. Obert Duval and Hedley formulas). Source: C. L. de Jongh
•	The compressive strength of a stope pillar is increased when later firmly confined by backfill because a triaxial condition is created in which s3 is increased 4 to 5 times (by Mohr's strength theory). Source: Donald Coates

Area: Subsidence
•	In block caving mines, it is typical that the cave is vertical until sloughing is initiated after which the angle of draw may approach 70 degrees from the horizontal, particularly at the end of a block. Source: Fleshman and Dale
•	Preliminary design of a block cave mine should assume a potential subsidence zone of 45-degrees from bottom of the lowest mining level. Although it is unlikely that actual subsidence will extend to this limit, there is a high probability that tension cracking will result in damage to underground structures (such as a shaft) developed within this zone. Source: Scott McIntosh
•	In hard rock mines employing backfill, any subsidence that may occur is always vertical and nothing will promote side sloughing of the cave (even drill and blast). Source: Jack de la Vergne

Area: Rockbursts
•	75% of rockbursts occur within 45 minutes after blasting. Source: Swanson and Sines
•	Seismic events may be the result of the reactivation of old faults by a new stress regime. By Mohr-Coulomb analysis, faults dipping at 30 degrees are the most susceptible; near vertical faults are the safest. Source: Asmis and Lee
•	In burst prone ground, top sills are advanced simultaneously in a chevron ('V') pattern. Outboard sills are advanced in the stress shadow of the leading sill with a lag distance of 24 feet. Source: Luc Beauchamp


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## alshangiti (22 أغسطس 2014)

: Compressed Air


Area: Air Intake


The area of the intake duct should be not less than ½ the area of the low-pressure cylinder of a two-stage reciprocating compressor. Lewis and Clark
Area: Air Line Losses


At 100 psi, a 6-inch diameter airline will carry 3,000 cfm one mile with a loss of approximately 12 psi. Franklin Matthias
At 100 psi, a 4-inch diameter airline will carry 1,000 cfm one mile with a loss of approximately 12 psi. Franklin Matthias
A line leak or cracked valve with an opening equivalent to 1/8-inch (3 mm) diameter will leak 25 cfm (42m3/min.) at 100 psig (7 bars). Lanny Pasternack
In a well-managed system, the air leaks should not exceed 15% of productive consumption. Lanny Pasternack
Many older mines waste as much as 70% of their compressed air capacity through leakage. Robert McKellar
Drilling requires a 25-psi air drop across the bit for cooling to which must be added the circulation loss for bailing of cuttings in the borehole at a velocity of 5,000 fpm, or more. Reed Tool
Except in South Africa, pneumatic drills are usually designed to operate at 90 psig (6.2 bars). Their drilling speed will be reduced by 30% at 70 psig (4.8 bars). Christopher Bise
A line oiler reduces the air pressure by 5 psi. Ingersoll-Rand
An exhaust muffler can increase the required air pressure by 5 psi, or more. Morris Medd
A constant speed compressor designed to be fed at 60 cycles (hertz) will operate at 50 cycles, but experience a reduction in capacity of about 17%. Jack de la Vergne
Area: Altitude


A constant speed compressor (or booster) underground will require 1% more horsepower for every 100m of depth below sea level. Atlas Copco
Auto-compression will increase the gage pressure of a column of air in a mineshaft by approximately 10% for each 3,000 feet of depth (11% for each 1,000m). Jack de la Vergne
The compressed air from a constant speed compressor will have 1% less capacity to do useful work for every 100m above sea level that it is located. Atlas Copco
Area: Cooling


A water-cooled after-cooler will require approximately 3 USGPM per 100 cfm of air compressed to 100 psig. Ingersoll-Rand
Area: Power


The horsepower required for a stationary single-stage electric compressor is approximately 28% that of its capacity, expressed in cfm (sea level at 125 psig). Lyman Scheel
The horsepower required for a portable diesel air compressor is approximately 33% that of its capacity, expressed in cfm (sea level at 125 psig). Franklin Matthias
To increase the output pressure of a two-stage compressor from 100 to 120 psig requires a 10% increase in horsepower (1% for each 2 psig). Ingersoll-Rand
Area: Receiver


The primary receiver capacity should be six times the compressor capacity per second of free air for automatic valve unloading. Atlas Copco
The difference between automatic valve unloading and loading pressure limits should not be less than 0.4 bar. Atlas Copco


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## alshangiti (9 يناير 2015)

[h=3]Cost Estimating[/h][h=4]Area: Budget Estimates[/h]

An allowance (such as 15%) should be specifically determined and added to the contractor's formal bid price for a mining project to account for contract clauses relating to unavoidable extra work, delays, ground conditions, over-break, grouting, de-watering, claims, and other unforeseen items. Jack de la Vergne
[h=4]Area: Cost of Estimating[/h]

A detailed estimate for routine, repetitive work (i.e. a long drive on a mine level) may cost as little as 0.5% of the project cost. On the other hand, it may cost up to 5% to adequately estimate projects involving specialized work, such as underground construction and equipment installation. Various
[h=4]Area: Cost of Feasibility Study[/h]

The cost of a detailed feasibility study will be in a range from 0.5% to 1.5% of the total estimated project cost. Frohling and Lewis
The cost of a detailed or "bankable" feasibility study is typically in the range of 2% to 5% of the project, if the costs of additional (in-fill) drilling, assaying, metallurgical testing, geotechnical investigations, etc. are added to the direct and indirect costs of the study itself. R. S. Frew
[h=4]Area: Engineering, Procurement, and Construction Management[/h]

The Engineering, Procurement, and Construction Management (EPCM) cost will be approximately 17% for surface and underground construction and 5% for underground development. Jack de la Vergne
[h=4]Area: Haulage[/h]

The economical tramming distance for a 5 cubic yard capacity LHD is 500 feet and will produce 500 tons per shift, for an 8-yard LHD, it is 800 feet and 800 tons per shift. Sandy Watson
Haulage costs for open pit are at least 40% of the total mining costs; therefore, proximity of the waste dumps to the rim of the pit is of great importance. Frank Kaeschager
[h=4]Area: Miscellaneous[/h]

The installed cost of a long conveyorway is approximately equal to the cost of driving the drift or decline in which it is to be placed. Jack de la Vergne
In a trackless mine operating around the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. John Gilbert
On average, for each cubic yard of concrete measured from the neat lines on drawings, approximately 110 Lbs. of reinforcing steel and 12 square feet of forms will be required. Jack de la Vergne
The overall advance rate of a trackless heading may be increased by 30% and the unit cost decreased by 15% when two headings become available. Bruce Lang
The cost to slash a trackless heading wider while it is being advanced is 80% of the cost of the heading itself, on a volumetric basis. Bruce Lang
[h=4]Area: Overbreak[/h]

The amount of over-break to be estimated against rock for a concrete pour will average approximately one foot in every applicable direction, more at brows, lips, and in bad ground. Jack de la Vergne
On average, for each one cubic yard of concrete measured from the neat lines on drawings, there will be two cubic yards required underground, due to over-break and waste. Jack de la Vergne


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## alshangiti (27 يناير 2018)

الاخ محمد معذرة على التاخير 
https://www.stantec.com/content/dam...14/Hard Rock Miner's Handbook Edition 5_3.pdf

HAND BOOK MINING


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