Design and planning parameters affected by ground-related problems:

Access Development:

Location / Orebody Characteristics: 

  • A typical ore body suitable for block caving is a porphyry-type deposit with disseminated mineralization having a large lateral and vertical extent.

  • It has been used in nickel, copper and molybdenum porphyries, hematite, asbestos, and diamondiferous deposits. It can be applied to steeply dipping vein deposits of sufficient width and in thick, flat-lying deposits. 

  • The rock strength can be from fairly weak to fairly strong, but in moderate dip tabular bodies, the total mass must have sufficient fractures in different orientations to allow the rock mass to break up under gravity into fragments small enough to pass through the drawpoints into the production drifts.

  • The lateral extent of the ore body must be large enough to ensure that a cave can be established. The horizontal footprint required to establish a cave will depend on the strength of the rock mass, but generally the minimum horizontal dimension of the mining area should be about 300 ft (90 m).

  • The height of the ore column should be sufficient to allow a reasonable productive life to the individual drawpoints and also to ensure a reasonable rate of return on the development and production costs.



  • If drifts on the extraction level are designed without respect to the direction of maximum principal stress, then overbreak can be intense in drift zones.  This is can arise through shear failure of the walls perpendicular to the maximum principal stress.  The overbreak can also occur at the intersection of drawpoint drifts and haulage drifts, and even in heavily supported drawpoints.

  • To minimize the overbreak and its consequences of reduced drift stability, the orientation of extraction and haulage drifts should be selected to suit the surrounding stress. 


Sublevel Interval: 

  • The haulage level is located under the production level  The distance between them depends to some extent on which system is used:

    • LHD system … the haulage level and the production level should be separated by a significant distance to provide adequate storage of production ore and to be able to reduce the number of raises to achieve the production target.  Raises should be inclined 60-70° to provide a good flow of rock and to enable the secondary fragmentation of the larger rock fragments as they drop down the ore pass.

    • Slusher system … the haulage level lies beneath the slusher loading hole, so the ore can be loaded directly into the haulage cars.  Long transfer raises tend to have been used, since each slusher drift should have its own connection to the haulage level.

    • With a grizzly level … the distance between it and the undercut depends on the pillar geometry needed to maintain grizzly level stability (commonly 5-8 m is used).

  • Typically the distance between the production level and undercut level is in the range of 12-18m.  This crown pillar offers protection to the production level. 

  • The Ertsberg Mine in Indonesia, has eliminated the need for a large separation of the production and haulage levels by using central storage bins and a distance of 10-15 m between the production and transfer levels. The production ore is fed through grizzlies into a central bin that discharges onto a 1.5 m wide conveyor belt that in turn discharges to a central crusher.


Production Drifts: 

  • In LHD systems, the production drifts are driven on even centres across the ore body.  The spacing between drawpoints will dictate the distance between production drifts.  The size of the drifts depends on the size of the production equipment selected.  The length of the unit is also an issue because the production unit has to manoeuvre into the drawpoint.  These drifts are roof bolted and will probably be concrete-lined or at least shotcreted.

  • Grizzly drifts have the following characteristics:

    • Connecting to the top of the transfer raises;

    • Serving as main accesses to the drawpoints;

    • Lined with concrete or timber to support against the stresses that may be generated during the undercutting and production phases;

    • Rockbolts may be used to support the pillar above the grizzly drift.



  • In a grizzly system, short transfer raises are usually driven between the production (grizzly) and haulage levels to allow for some storage of production ore for loading the haulage units.

  • Draw raises are driven from the grizzly (production) level to the undercut level.  Two raises are drilled on opposite sides and at right angles (not less than 60°) to the grizzly drift and located at the top of the transfer raise.  The brow at the intersection of the draw raise and the grizzly drift is supported with posts and beams to help maintain its size.  The function of the brow in the draw raise is to hold the broken rock inside the draw point and avoid the muck pile surging out into the grizzly drift.  In Grizzly systems, the number of grizzlies that can be served by one transfer raise will depend on the distance between levels and the grizzly chamber spacing.

  • The transfer raises are usually lined with timber or concrete to ensure that their life exceeds the life of the production from above.

  • Chutes are commonly installed at the bottom of the transfer raises, although other methods may be used. Grizzly rails are installed at the top of the transfer raise in order to size the material passing down the raise.

  • The cross sectional size will be determined by the amount of ore storage required within the constraints of the ground conditions. Some recommend a cross section size of four times the diameter associated with the 15 % passing of the production fragmentation.


Undercut Layout: 

  • The design and sequence will be dependent on which of the three block caving systems is used and on the configuration of the ore body.  The haulage and the undercut level are fairly standard whereas the production level will be substantially different.

  • The undercut drifts are usually driven directly above the production drifts. 

  • The undercut drifts connect the top of the upper draw raises.  Pillars are left between undercut drifts that will be blasted later in the operation to generate the caving.

  • The size of the undercut drifts is a function of the size of the equipment that will drill and blast the undercut of the block. Undercut drifts are temporary. There is therefore no need to change the size of the undercut drifts compared to the production drifts, although sometimes the equipment to drill the blast holes needs more space than the LHD. For the same reasoning, these drifts are not necessarily supported, unless ground conditions require some temporary support.

  • The design and operation of the undercut is the key mechanism for initiating the cave, (Calder, 2000).  As an example: The Palabora Mine undercut design has the following features:

    • An advanced undercut is used, whereby the undercut is developed ahead of opening the draw bells to provide a stress shadow to protect the production level.

    • A narrow undercut 4 m high is excavated. This is considered sufficient to initiate cave while minimizing the amount of swell material that has to be removed during excavation using the advanced undercut

    • An inclined face is made over the major apex creating a chevron shape to the undercut. This is to facilitate undercut initiation following the construction of the draw bell by promoting the flow of blasted undercut material into the bell.



  • The undercut level in traditional panel caving is located 18 m above the production level (Jofre, 2000). Undercut blast holes, approximately 50 mm in diameter, are drilled from drives 3.6m wide by 3.6m high, forming a void where panel caving may be induced.

  • At the El Teniente Mine the height of the undercut level is considered to be the main parameter in the undercut design.  The design of the undercut has evolved over the years.  Even though many workers have recommended a height of 10.6 m, there is still the trend of using a height of 3.6 m around the world.

  • Reducing the height of the undercut can cause problems.  Some of these have been seen at the El Teniente Mine, such as reduced likelihood of achieving caving, less volume provided for rock expansion and also, production of a coarse initial fragmentation (more secondary breakage is therefore needed, which can damage the drawpoint, brow, support and flow).

  • The design of the undercut in the Esmeralda sector within the El Teniente mine, is based on tunnels driven 3.6m x 3.6m developed parallel to each other on 15m centres.  The excavation of the undercut is achieved by blasting three or four fan holes drilled into the sidewalls 7-10m in length. 



Support Requirements: 

  • As an example, the following reviews the tunnel support requirements at the Premier Mine (Bartlett and Nesbitt 2000):

    • Tunnel development is generally carried out in good rock conditions.

    • Resin grouted 1.8 m long rockbolts are installed on a 1m2 spacing within 5 meters of the advancing tunnel face. The fracture zone associated with tunnel development is usually less than 0.5 m and at least 1.0 m of the 1.8 m long rock bolts are resin grouted into solid rock.  Problems arise only once tunnel development starts to be affected by undercut abutment stresses.

    • The support at the undercut level is kept to the minimum commensurate with the safety of workers and equipment.  Tunnels may be affected by abutment stress up to 20 m from the front cave.  In addition to rock bolts, tunnels are supported with chain link wire mesh and steel tendon straps.  Shotcrete is applied to the height of 2 meters on the sidewalls to prevent LHD damage to the mesh and tendon straps.

    • Intersections of tunnels are supported with longer roof bolts and corners are strapped with steel cable trusses, covered with shotcrete.  This level of support is usually adequate unless water and/or high stresses affect the area.

    • The support on the production level is a function of the abutment stresses produced by the movement of the undercut front (Bartlett, 2000).  Because of this, some rules have been established to minimize the effect of induced stresses on the production level:

      • If the stress increase exceeds 5% of the uniaxial compressive strength of the rock then shear movement on existing joints and fractures tends to occur.  If the area has been supported with good inter-bolt support, such as shotcrete reinforced with mesh and tendon straps, then  this movement is effectively constrained.

      • If the stress change exceeds 20 % of the uniaxial compressive strength of the rock this is usually enough to cause the propagation of the fracture zone around the tunnel by failure in shear of the rock.  The result is tunnel convergence in rigid concrete or concrete linings, leading to limited failure of the lining.  Overall, rock bolts and reinforced lining provides sufficient constraint to keep displacements within stable limits of less than 50 millimetres.

      • If the stress change exceeds 50 % of the uniaxial compressive strength of the rock then this is usually enough to induce extensive failure in shear of the rock around the excavation.  Large tunnel convergence in excess of 200 mm may occur which sets up the destruction of the rigid concrete, leaving the rock between the rock bolts without constraint .  Finally the tunnel crushes.


Development (Undercutting) Sequence:

  • Drilling of the longholes for the undercut and sometimes for the draw bells is done from the undercut level.  The original design of the LHD block caving system incorporated the drilling of the draw bells from the undercut level to ensure proper blasting sequence.  Today, the draw bells are developed entirely from the production level and the blasting is started by drilling a raise in the centre of the draw bells. 


  • Blasting the draw bell immediately precedes blasting of the undercut. In this way the minimum amount of ground ahead of the front cave is opened and more support is offered to the weight that precedes the cave. Longhole blasting is done by two to three fans at a time. Using this method it is reasonably easy to check for unblasted pillars. The front cave should be advanced at a certain angle to the major workings to minimize the amount of stress transferred to the production level.


  • In any cave mining system the block to be caved must be undercut to induce caving.  Access must also be provided for extraction units to load ore from the base of the cave and transport it to the passes or crushers.  The excavations associated with block caving can be developed in at least 3 different sequences.  The sequence determines the extraction ratio on the production level prior to the level being subjected to induced undercut abutment stresses.  The mining sequences are termed post undercutting, advance undercutting and pre undercutting.


Post undercutting 

  • In this sequence all of the development on both the extraction and the production levels is completed before undercutting starts.  The major advantage is that the undercut level does not have to be developed and equipped like the production level does.  Blasted and caved ore is removed quickly to avoid compaction.  When using this method, tunnels are subject to high and variable stress in terms of both magnitude and direction.  This might result in extensive rock and support damage, depending on rock mass strength, mining depth and support effectiveness.

  • Undercutting in traditional panel caving requires the complete development and construction of the undercut and production levels.  Draw bells are opened before the passage of undercut blasting.

  • Unfortunately this sequence exposes the production level to high values of abutment stress from the advancing undercut front, inducing potential damage in the production level crown pillars.


Advance undercutting 

  • In this sequence the objective is to achieve as low an extraction ratio on the production level as possible, prior to excavations being subject to high undercut abutment stress.  The amount of development on the production level is determined by rock mass strength, stress levels and the minimum development needed to achieve the production target after undercutting, so that ore can be mined before compaction occurs and stresses are re-established.

  • The major advantage is that rock and support damage is constrained to acceptable levels as a result of a low extraction ratio on the production level.  The major disadvantage is that the undercut level should be implemented as a production level. 

  • Support is carried out in three phases, and logistics are also complicated.  Development, through opening and support must be carried out quickly to avoid compaction.  

  • Advance undercutting has been successfully implemented at the Premier, Esmeralda and Palabora Mines.



  • In this sequence the undercutting is completed before any development is carried out on the production level.  The major advantage of this method is that production level development is carried out in de-stressed rock.  Rock and support damage are minimized and support is completed in a single phase.  The major disadvantage is that production level development must be carried out very quickly once undercutting has been completed.  Experience shows that compaction may occur after 6 month at a mining depth of 630m below surface and within 3 months at a mining depth of 730 m below surface.  The development should be very carefully planned to ensure that production would start before compaction takes place.

  • The rate of development must be balanced with the production rate.  Generally, if the number of new tons caved during a year is equal to the number of tons drawn, then the productive capacity should remain constant.  If the tonnage above the individual drawpoints varies significantly, then some adjustments must be made so that the number of available drawpoints is adequate to meet the production requirements.

  • The amount of lead-time required to prepare the drawpoints must be carefully calculated.  The lead-time required will depend on the available manpower and the efficiency of the manpower and equipment being used.  Generally, about three years will be required to prepare a new production area  from the time the initial access is started until the undercut longholes are blasted.  Undercut drifts should be completed about six months to one year before undercut blasting to allow adequate time to complete the longholing. Longholing should precede blasting by about six months.



Caving Front Orientation: 

  • The shape of the cave front can modify the stresses in the cave back.  Numerical modelling can be a useful tool in helping to determine the stress pattern associated with several possible mining sequences.

  • The orientation of the cave front/back with respect to the joint sets and direction of principal stress can have a significant effect on primary fragmentation.

  • The abutment stress generated at the caving front is the main source of the damage to the drifts located at the lower levels.  Stagnant cave fronts should be avoided since this causes column or point loading on the apexes.  Large irregularities in the horizontal geometry of the cave line cause an increase in stress concentration, resulting in serious damage on the undercut front.  It is therefore imperative that all large irregularities are kept to a minimum.


Fragmentation Control: 

  • The central design parameter of a block caving operation is fragmentation.  Two types of fragmentation are involved in caving methods - primary and secondary - both affect the overall design, draw control, productivity and cost of the mining system.


  • Caving results in primary fragmentation, which can be defined as the size distribution of the particles that separate from the cave back and the draw column (Laubscher, 1990). Primary fragmentation is produced due to stresses acting in the cave back, primarily along favourable oriented joint sets with some shearing of intact rock. The orientation of the cave front or back with respect to the joint sets and direction of the principal stress has a significant effect on the primary fragmentation.  The size of the potential rock blocks is based on the adjusted joint spacing.

  • Secondary fragmentation is the reduction in size of the original particles that enter the draw column as they move down through the draw column.



Muck Flow / Cave Control - Cavability: 

  • After determining that the size of the ore body is large enough to justify the block caving choice, then the cavability must be determined.  RQD can be the first indicator for determining cavability, but is does not provide: fracture orientations; fragmentation intensity; or the overall characteristics off all fracture sets.

  • Rock mass classification systems can be used to define cavability. These systems incorporate information such as joint spacing, joint conditions, intact rock strength, joint filling and other factors. Laubscher (2001) defines the Mining Rock Mass Rating (MRMR), which is the Rock Mass Rating (RMR) adjusted by weathering (75%), orientation of the cave front (70%), induced stress (120%-76%) and blasting (80%). Laubscher relates MRMR and the hydraulic radius (area/perimeter) that represents the size of the excavation to define the cavability of an ore body. An empirical graph that represents the world wide caving and stability relationships is well established to help determine if a block will cave or not.                                         

  • There are two different caving mechanisms that act in a block caving mine.  These have been defined as stress and subsidence caving. 

    • Stress caving occurs in a virgin rock mass when the stresses in the cave back exceed the strength of the rock mass.  Caving stops when a stable arch is reached, then the undercut must be increased in size, or boundary weakening must be undertaken to induce further caving (Panek, 1981). 

    • Subsidence caving occurs when adjacent mining has removed the lateral restraint on the block being caved.  This can result in rapid propagation of the cave, with limited bulking (Laubscher, 2001)

  • Stress analysis shows that excavating a tabular opening (undercut level) into a compressive stress field causes redistribution of the pre-existing stresses surrounding the opening in such a way as to:

    • relieve the vertical compressive stresses within a dome-shaped zone (R) above and below the slot

    • transfer these stresses, causing higher than normal compressive stresses in the abutment zone (C) around the end of the slot.

    • in response to relief of the compressive stress, the ground below the opening moves up (v) and the ground above the opening moves down (v).


  • If the ratio of  the horizontal to vertical (lithostatic) stress exceeds 0.5, then there are no tensile stresses in R.  If the ratio is less than 0.5, then the tensile stress may cause some initial sloughing, up to a dome shape.  For caving to continue upward, above the dome or arch, a breakdown must be initiated - produced by compressive or shearing failure - because tensile stresses are absent. 

    • One approach is to analyse the rock mass as a continuum, then evaluate the planes of failure produced by an excess of shearing stress over the available shear strength.

    • Another approach is to analyze the displacement of the blocks with respect to their neighbours under the effect of compressive stress.  The arch will progressively sag, opening up spaces between the key blocks and losing its structural integrity. 

    • Another mode of breakdown is the gravity fall or the extruding of small fragments from sheared, shattered, or crushed material that intersect the arch.  Each small fragment that falls out removes some constraint to the movement of its neighbour.  This progressive dribbling or ravelling is commonly observed in underground excavations.

  • In summary, the cavability of an ore body is strongly influenced by geologic factors, such as fracture intensity, fracture attitude, faults, rock mass strength properties, post-ore dykes and the stress field.  These can be examined during the cavability portion of the design process by determining the domains that are present, both in the ore body and the surrounding rock. Geologic mapping and diamond drill core logging, fracture maps and rock mass strength tests will thus help to define the cavability of an ore body.


Induced Caving Stress:

  • High abutment stresses are an inevitable part of cave mining.  At shallow mining depths and in competent rock the induced stresses are not sufficient to cause large-scale damage.  With increasing depths alternate mining sequences must be implemented to avoid tunnel and support damage.

  • Data from the El Teniente Mine demonstrated how the abutment stress could be three times the value of vertical stress.  The abutment stress acts in a range of 60-120 m from the undercut front.

  • Important parameters affecting the production level prior to undercutting, such as the extraction ratio, the shape of the undercut face, the rate of undercut advance and leads and lags between tunnels must be rigidly controlled. (Bartlett and Nesbitt, 2000)

  • Abutment stress produces extensive damage to rigid shotcrete linings.  A phased approach to support installation limits the degree of damage.  Final rigid linings and concrete footwalls are only installed once the undercut abutment stresses have passed over the area.


Particle Flow:

  • An understanding of the flow of granular material (e.g. rock, sand, grain, prills) as with SLC is the first step toward understanding block caving draw.  If a supported column is taken and gravity flow is then allowed through an aperture at the bottom of the column, an ellipsoidal shaped zone of disturbance of the in place material will form and grow with the continued extraction of material through the aperture.  The volume of this zone of disturbance or loosening is about 15 times the volume of material extracted at any given time.

  • Over time, a zone of loosening develops.  The ellipsoid of loosening represents a future ellipsoid of extraction which will equate to the volume of material extracted by that future time. Extraction involves the movement of the particles located vertically over the drawhole directly downward, while material outside this direct vertical volume moves radially inward and down toward the draw point. (Richardson, 1981)

  • The ratio of the height over the width of the loosening area increases as ore is withdrawn from the draw column (Kvapil, 1965).  Therefore a vertical cylinder might be used to better represent a draw column in block caving where there are no bounds.  The width of the cylinder is a critical dimension for the determination of draw point spacing in relation with block caving mine design.

  • The overall control of the draw column width is the mobility or lack of resistance to flow motion of the material being drawn.  Material with a high degree of mobility will show a thin draw column, which will not support a very large arc without failure.  The most universal parameter to measure or relate the mobility of a rock mass is the particle size.  Bin experimentation has shown many relations between the particles size and the potential width of the draw column. 

  • The drawing of the caved ore reveals a funnel-shaped zone (F1) in which the particles move steadily toward the draw point.  Near the bottom, the sides of the funnel are inclined about 70 degrees to the horizontal; with height, the inclination tends to be 90 degrees.  Particles outside the active draw zone move down and toward the funnel; the boundary of this zone is called F2.  The loosening due to particle migration in the second funnel permits strain relief SR to occur in the surrounding rock.  The San Manuel Mine projects an inclination of the total zone of loosening to be represented by an angle of 60 degrees.

  • Besides the rock fragment size there are many others factors which affect mobility and width of draw, as follows:

    • Shape of the rock fragments

    • Angularity of the fragments

    • Roughness of fragment surfaces

    • Stickiness

    • Fines within the coarse fragments

    • Depth or gravity pressure on the rock column

    • Strength of the fragments

    • Moisture content


Draw Points / Draw Bells: 

  • Draw points are drifts driven horizontally from the production drifts.  Usually they are driven at an angle to the production drift to facilitate the entry of LHDs to the draw point.  The distance from the brow of the draw point to the opposite rib of the production drift must be sufficient for LHD to have entry to the muck pile with near zero articulation. 

  • The height of the brow should be enough to allow the bucket of the LHD to raise its load but not so high that the ore will flood on the production drift.

  • Draw points are lined with concrete to maintain the size and avoid major deformations.  Steel arches (3-6) as well as rock bolts (5-10) are used in the brow to ensure the life of the draw points.

  • Two adjacent draw points are connected by a crosscut called a draw bell cone.  The draw bell is used to create the draw zones feeding broken material to two draw points.

  • For example, the production level at the Premier mine had been developed as an offset herringbone layout. This layout was chosen after several options had been considered and modelled for ease of operation, structural strength, and the use of electric LHDs.  The production tunnels are 4 by 4.2 m in size, and are spaced at 30m centres, the draw bells are at 15m centres.  The draw bells were planned to be 13 m long to allow for drawpoint wear of 1m at either end. (Bartlett, 1992)

  • The draw point for the grizzly and slusher systems are driven at right angles to the grizzly or slusher drift. They are also inclined to the horizontal, the angles varies nearly 90° to 45°.  The cross section should be large enough to allow the fragments to pass through the opening with minimum secondary breakage.

  • After driving the inclined section of the draw point, a vertical section may be driven upwards to the elevation of the undercut level.  This allows the undercut level to be raised further above the production level to ensure an adequate pillar between the cave area and the production level.

  • A concrete liner will allow a better flow of the broken rock and will retain the size of the draw point better than unlined rock.  If repairs are required because of the excessive erosion, then it will be easier accomplished if the draw point was originally lined.  It is very important to maintain the size of the draw points to prevent flooding of the production drift with broken rock.                           

  • In an LHD system, the draw point entry must be nearly horizontal to allow the entry of the production unit. The connection between the production level and the undercut level is usually a large draw bell created by drilling and blasting.  The draw bell allows for larger pieces of rock to be moved by the LHD.

  • The brow of the draw point should be high enough to allow the LHD bucket to lift a load, but not so high as to allow broken rock to flood out into the production drift where the movement of the LHD could be restricted.

  • Draw point spacing is important to insure a good recovery of the ore.  If the draw point spacing is too far apart, then good ore will be lost or dilution may become so great that the ore becomes uneconomic.  If the draw point spacing is too close together, then the mining costs are greater than they should be.  Each draw point has a certain area of influence on broken rock above.  The draw point spacing should be spaced that these areas of influence will overlap slightly to insure that the total column is moving downward as the draw progresses.

  • White (1979) has constructed a regression curve based on data that can be used to approximate the draw area based on median fragment size.  For example if the median fragment size is 0.8 m, then the draw area would be equal to 131 m2 (y=152x0.8+9.8=131.4m2)   


Mine Median Fragment Size (m) Draw point spacing (m2)
Ertsberg 0.7 236
El Teniente 0.7 224
Grace 0.8 165
Henderson 0.5 148
Creighton 0.5 110
Climax 0.9 106
Urad 0.6 80
Thetford 0.5 58
Mather 0.2 32
San Manuel 0.4 24
Fragment size & Drawpoint spacing in various mines  (White, 1979)


  • Drawpoint spacing for grizzly and slusher layouts reflect the spacing of the draw zones because of the close spacing of the draw points (Laubscher, 2001).  However, in the case of LHD layouts with a nominal draw point spacing of 15m, the draw zone spacing can vary across the major apex (pillar) from 18 to 24 m, depending on the length of the draw bell.  There is a trade off between optimizing the ore recovery through the use of a short draw bell and the desired usage of the LHD.


  • Sand-model tests have shown that there is a relationship between the spacing of drawpoints and the interaction of draw zones.  Widely spaced draw points develop isolated draw zones, with diameters defined by the fragmentation. When draw points are spaced at 1.5 times the diameter of the isolated draw zone (IDZ) then interaction among rock fragments occurs.  The interaction improves as the draw point spacing is decreased, the question is whether the interactive draw theory can be wholly applied in coarse material, where arches of more than 20m have been observed.  Low-friction material could flow greater distances when under high overburden load, and this could mean a wider draw zone spacing.

  • The sand model results (Kvapil, 1965) have been confirmed by observation of the fine material extracted during cave mining and by the behaviour of materials in bins.

  • Vertical stress increases just over the brow of the drawpoint.  This effect demonstrates that particles tend to go to the centre of the draw zone, exhibiting a horizontal component of the total particle’s movement.

  • Finally, according to the design criteria of the Premier Mine, the drawpoint spacing, largely a function of fragmentation, must be addressed to reduce hang ups, maintain production levels, and avoid damaging stress levels.


Rate of Draw / Oversize: 

  • The caved material should be drawn rapidly enough to avoid packing in the new undercut and permit the caving to proceed upward through the rock mass.  The rate of drawing broken rock may vary, from 152 mm to 1.2 m per day.  Material from newly undercut pillars should be pulled immediately following an undercut blast to be certain that no pillars are left behind.  Oversize chunks must be blasted to pass through the draw points.

  • In recent years the rate of draw also has been recognized to be potentially constrained by the risk of rockbursts.  The El Teniente mine limits the draw rate in the first 30% of the draw column to be less than 0.3 tons/m2 per day to avoid potential rockbursting. 

  • Observed seismicity in block caving has the following features:

    • The seismic activity in magnitude and number of events is strongly related to mining activity;

    • The seismic activity in block caving  can be related to:

      • initial caving;

      • the fragmentation process;

      • the in situ stress field.


  • The rock mass surrounding the extraction openings is subjected to four stress cycles in all caving situations. The cycles are the following: 

    • Adjustment of the rock mass to the opening

    • Abutment stresses development ahead of the undercut, and re-distribution of stresses around the caving area

    • Uplift after the undercut is complete with the removal of the vertical stresses

    • Vertical loading on the apexes from point loads and an increasing column of caved material.

  • The vertical stress on the extraction level pillars is related to the draw height and draw area. Cave model studies (Heslop and Laubscher, 1981) have shown that for dynamic conditions, with height to base ratios of 1:1, 2:1, 3:1, 4:1 and 5:1, the average vertical stress on the pillars is approximately 53%, 30%, 22%, 17% and 14 % respectively of the mass of the cave ground.


Ratio vertical :
Base: 60m
Base 100m
Base: 200m
1:1 0.8 1.32 2.64
2:1 0.9 1.5 3.0
3:1 0.99 1.65 3.3
4:1 1.02 1.7 3.4
5:1 1.05 1.75 3.5
(Heslop & Laubscher, 1981)


  • The increase in vertical stress with static conditions will be approximately 10% of the rock mass weight and stresses in the centre of the area under draw will be higher than at the sides.  Hang-up areas with well-developed arches (large wedges supporting columns of caved material and consolidated material) will concentrate the vertical stress.  In the case of a 200m diameter extraction level with 5:1 ratio point loads of 100 MPa or higher could be expected (Heslop & Laubscher, 1981).


Dilution – Recovery:

  • It is not possible to predetermine the exact dilution, but some estimate can be made based on experience at other similar mines.  The dilution typically can vary from 10 to 25% of the total ore drawn.  The amount of dilution that can be accepted often is a function of ore grades, grade of the diluting material, costs, and metal prices. 

  • Well-controlled draw is essential to minimize effects of dilution.  A dilution zone that breaks into the same size as the ore or larger is best for minimum dilution.  Where waste is finer than ore, greater amounts of dilution can be expected but should be controlled with good draw practices.

  • The recovery of ore and amount of dilution are important factors in the success of a caving system, and there is no known method of predetermining the end results, although experience at other mines could be a good indicator.  (Tobie and Julin, 1982).


Mine % of Ore recovered % of Dilution
Cornwall 100 17
Grace 85 20
Mather 67 10
Thetford 100 20
Creighton 95 15
Climax 92.5 15
San Manuel 101.5 12
Urad 101 15
De Beers 100 20
El Teniente 91 10-20
Recovery & dilution at various mines (Tobie and Julin, 1982)


  • Laubscher (2001) defines the height of interaction as the main parameter to predict the percentage at which the first dilution would appear.  The height of interaction (HIZ) is the height where interchange of material is happening between drawpoints, above the HIZ the movement is descendent uniformly.  The ratio between the HIZ and the height of the draw column represents the percentage at which the first dilution appears.

  • Laubscher (2001) estimates HIZ based on fragmentation, draw point spacing, draw control and the swell of in-situ rock. 


Water Inflow (Mud Rushes): 

  • The main causes of mud rushes are:

    • Underground workings daylighting to surface

    • Poor water control and drainage management system design in a caving operation will result in groundwater and rainwater entering the cave zone above the current workings.

    • Poor waste management practice, e.g. dumping of tailings and slimes in old open pits and siting of tailings dams above caving operations.

    • Poor draw control practices, causing ingress of mud or water by drawing the water capping above the ore block.

  • Safe, effective mud rush control is achieved by applying the following principles.

    • Effective drainage of the mine area, both in surface and underground workings at the muck pile

    • Implementation of good draw control standards to prevent drawing of the mine waste capping which may contain mud.



  • Surface subsidence will always occur with Block Caving.  The subsidence will depend of several factors such as depth of mining below surface, strength of ore, and rate of draw among others.

  • The first sign of subsidence is a circular hole or funnel appearing somewhere within the boundaries of the undercut area.  Eventually the whole caving area will settle, and an escarpment will form.  Escarpments found in mines vary from 45° to 80°, depending on the factors mentioned earlier.  Typically a 60° line is drawn from the undercut level to the surface to estimate the potential area where cracking might occur. No structures should be placed inside that area because there is a high probability to affect them due to ground strain and tilt.