Vertical crater retreat
From Queen's University Mine Design Wiki
- This article is about the orebody requirements and developmental steps involved in Vertical Crater Retreat planning and mining.
Vertical crater retreat (VCR), also known as Vertical retreat mining, is an open stoping, bottom-up mining method that involves vertically drilling large-diameter holes into the orebody from the top, and then blasting horizontal slices of the orebody into an undercut. A system of primary and secondary stopes is often used in VCR mining, where primary stopes are mined in the first stage and then backfilled with cemented fill to provide wall support for the blasting of successive stopes .
Similar to Sublevel open stoping and Blasthole stoping methods, VCR mining is used for steeply-dipping (>45º), or both vertically and horizontally large orebodies with competent ore and waste rock strength. It differs from other open stoping methods in that it is a bottom-up method, as opposed to a left-to-right method, and it does not require the excavation of sublevel drifts before blasting and mucking can take place. The thickness of one horizontal slice varies between 2 and 5 meters in height.
- 1 Orebody Characteristics
- 2 Advantages
- 3 Disadvantages
- 4 Overall Mining Process
- 5 Mining
- 6 References
VCR is a suitable mining method for orebodies that exhibit the following characteristics:
' 'Orebody Dimensions
|Ore and Rock Strength||
- Wall Support: VCR stoping shares some great features with sublevel open and shrinkage stoping. Good wall support is offered during the VCR stoping phase, using shrinkage 
Overall Mining Process
The first characteristics to evaluate are the size, dip and plunge of the orebody, which is important because the installations of draw points are essential to the gravity flow of the blasted ore for collection. The second element to assess is the shape and consistency of the orebody. Two horizontal drifts are required before mining can take place, which are to have a very large vertical separation. The distance between the two drifts depends on the consistency of the ore, the drilling accuracy, accessibility, and competency of the hanging wall. These drifts are cut inside the orebdoy in order to minimize developmental costs. The next step is assessing the blasting characteristics of the rock, which will help to determine the drilling pattern and stope sizing of the mine. These tests can be done on similar ore blocks, or simply theoretically. Early consideration of equipment selection can be done at this point, as they will be based on stope and block size, as well as production requirements, and most importantly availability .
Development of Infrastructure
Once the essential planning is complete, development of the pre-mining infrastructure begins. All working drifts are reinforced with the necessary ground support. The pre-mining development for VCR mining includes:
- Haulage drift along the orebody, at the drawpoint level
- Drawpoint loading arrangement below the stope
- Undercut of the stope
- Overcut - as access for drilling and blast charging 
Holes are drilled vertically from the top drift through to the bottom drift. Holes are charged such that blasting of horizontal slices of the orebody occurs, progressing from the bottom drift to the top drift. In any new region of the mine, the ore zone is assessed as soon as possible, so that ore data can be collected and compared with the original estimates calculated by mine engineers. This offers the engineers a chance to analyze the data and make any required modifications before the following stope layout is planned. Extraction of the blasted material can now occur as fast as the system is designed, however just enough broken ore is mucked from the stope to create the required volume of space for successive blasts. Blasted ore is collected at the drawpoints using LHD vehicles, and then transported to orepasses, sometimes to be crushed, before it is transported to surface for processing. Upon completion of the ore extraction, the stope is often backfilled from the top drift, providing rock stability for upcoming blasts. This process is repeated until the orebody is mined.
Stope Development and Access
Stope layouts, like all components in the mining process, are continually modified and updated based on discoveries that occur during mining. Based on the drilling exploration program, an outline of the orebody is developed, which influences the layout, size and shape of the stopes. Geologic data is the primary information used for developing stope layouts. Plan views of the orebody are developed, and then stope plans begin, with the goal of minimizing required development and optimizing the drilling and blasting within the stope.
Based on the ground conditions, orientation and size of the stopes, pillars may be left in the top sill. The distance from the top sill to the ore itself is designed to be very small, to maximize ore extraction while minimizing development. The drift is arched, and is typically 9 x 9 ft (2.7 x 2.7m) or 10 by 10 ft (3 x 3m), based on the size of the LHD equipment operating in the area.
The development of the top sill consists of driving an access into the orebody, and then cutting out the ore in that region, along with any waste also inhabiting the zone. Once complete, ground support is implemented so that working crews can maneuver safely within. This is the platform from which drilling occurs, going down to the bottom sill.
The main haulageway is usually driven along the strike of the orebody, with approximately 40 to 50 feet separating it from the bottom sill. Scrams are then driven into the sill every 30 to 50 feet along the strike, with the actual spacing dependent on the ground conditions. The smaller the spacing, the higher the potential extraction ratio (without the required use of remote controlled mucking). However, a higher extraction ratio means that smaller pillars are left behind, which raise potential stability issues.
In some cases, the stopes are designed to be stacked. If this is the case, mine engineers will often make the top sill from one stope into the bottom sill for the next stope. Access to the orebody from the top sill can thus be utilized to gain access to the ore during the extraction from the next sill.
Extraction of the ore is done from drawpoints at the bottom sill. The design of successive stopes is very important to ensure that equipment can get access to the blasted ore after blasts. In some cases, the ore may extend below the bottom sill, in which case a cut of approximately 12 ft can be taken out of the sill. This ore will be extracted using the system in place, and waste material will be brought in to replace the cut made in the sill.
Productivity of a stope in a VCR mining is directly related to the stope size. Large stopes have high tonnages that can support high mucking rates for long periods of time, resulting in higher productivity rates.
The homestake mine in South Dakota is a VCR mining operation. Their switch from cut and fill to VCR mining has been a major factor in their increased productivity since the 1970s. The next best method employed at Homestake is mechanized cut-and-fill (MCF), which has proven to be at a 48% cost disadvantage to VCR, in terms of direct mining costs. The overall Mine Department productivity has almost doubled, from 5.94 tons/employee shift in 1979 to 10.38 tons/employee shift in 1988. The cost advantage is primarily contributed towards higher productivity and less required ground control. The productivity of a mining method will vary based on the dimensions of the orebody, however if the orebody meets the requirements laid out Orebody Requirements section, then VCR mining will provide high productivity rates, when compared to other methods.
During the stage of mine development, the preferred drilling equipment is the diesel Jumbo, a one or two-boom pneumatic hammer drill. With one worker at the control panel, a Jumbo drill can complete a pattern of 60 blast holes, each at a depth of 4 meters, in just a few hours 
Longhole Production Drill Rigs are used for the production blastholes. These machines are equipped with a powerful hydraulic rock drill, as well as a carousel storage for extension rods. In addition, drilling can be performed by remote controls, so that the operator may remain in a safe position . Down-the-Hole Drills can also be used for the blastholes, which operate by breaking the hard rock into small flakes and then blowing them clear using an exhaust system.
Load-Haul-Dump (LHD) Vehicles are the most commonly used mucking equipment in VCR mining. These four-wheel drive, rubber-tired machines are either electric or diesel powered and maintain a low profile. They are centre-articulated front end loaders and are common in various mining methods, as they provide in-stope mobility. These machines operate using buckets that range in size from 0.5 yd3 (0.38m3), for "micro" scoops, to 14 yd3 (10.7m3) for large scoops. A typical unit with a 4 to 6m3 bucket size, transporting material over an average distance of 150m, can move between 550 to 800 tonnes of material per hour.
LHD vehicles however have high tire wear costs, due to their constant operation on rough ground surfaces and movement over irregularly-shaped rock fragments. These tire wear costs can range from $0.75 to $3.00/tire per hour, therefore total unit costs have been calculated to range between $3.00 and $12.00/hour. In addition, mines incur a large capital expenditure when purchasing new tires for these machines, with prices per tire ranging from $4,000 to $6,000 .
Some VCR mining operations use CAVO Muckers to transport material from the drawpoints to the crushers or orepasses. Cavo Muckers are made by Atlas Copco, and they are maneuvered by an operator that stands on the side. They scoop material and dump it into their bucket and when it is full, the material is dumped out of the back of the machine. CAVO Muckers are typically used in remote, smaller stopes, as well as in hard-to-ventilate areas .
Boreholes are developed to join the scram system to the loading chute of a haulage drift, between 150 to 300 feet below. Material is moved from the drawpoints to the borehole with LHD equipment. Rock flow at the bottom of the borehole is controlled using a power chute. In many cases of VCR mining, the rock traveling down the borehole is met at the loading chute by rail haulage cars, which then transport the ore to crushers, and eventually to surface. Attention must be taken when planning installation of the power chute, as it can often cause delays in production if its implementation is not properly coordinated.
The main haulage drift is usually track, and is implemented 30 to 40 feet from the bottom sill along the strike of the orebody. Scrams are often spaced as closely together as possible, which maximizes extraction because remote mucking is most often not feasible.
Drilling is done from the top sill all the way down to the bottom sill. The cross-sectional dimensions of the top sill must be 11ft (3.4m) high and 15ft (4.5m) wide in order for the drill mast to achieve clearance and maneuverability: . The most common drill hole diameter used in VCR mining is 165mm, however 140mm diameter holes have been used as well as 205mm diameter holes, in some rare cases. The benefit of using 165mm diameter drill holes is that it allows for a simple 4 x 4m drill hole spacing. Drill holes can be up to 60m (200 ft) in length.
Drill pattern configurations are modified and adapted for every stope, and they differ from mine to mine. General guidelines dictate that an 8 by 8 ft (2.4 by 2.4m) pattern is used, in order to minimize breaking in the handing wall and drill footage. However, this can vary up to 10 by 10 ft, based on ore recovery and changing geometries of the orebody between the sills.
Holes drilled from the top sill right down to the bottom sill are called “breakout” holes. As drill holes are so lengthy, accuracy is very important. Drill accuracy is determined by the drill pattern laid out in the top sill. Often times mines will check the bottom sill for proper alignment before drilling of the hole is complete. This is done to determine if there is a large amount of deviation from the planned pattern, in which case additional holes would need to be drilled to correctly cover the stope.
Once everything is set up, drilling can proceed at a fairly rapid pace. Many drill crews can average 200ft (60m) in a shift.
Before the drill holes are loaded and blasted, they are first measured. Typically this is done using a piece of wood approximately 3 ft long, attached to the bottom of a rope and lowered into the hole. Once it hits the bottom surface it is pulled up until the stick is horizontally flush with the bottom of the drill hole. If the bottom of the hole has not broken through the rock, the stick may be substituted for a steel weight.
After all of the holes in the stope design have been measured, the amount of void space between the bottom of the drill hole and the surface below must be estimated, in order to account for the swell of the blasted material. The next step is to determine the position at the bottom of the blast hole that should be blocked, for explosives to be placed on top. The angle of the blast hole dictates generally how high the block distance is above the bottom of the hole. Sample blocking height values, based on hole angles are displayed in the following table.
Blocking Height (Above Bottom of Hole)
| 80 - 90
| 57 - 79
| 50 - 56
| Less than 50
Once the blocking values are determined, loading of the hole can begin. A wooden wedge is tied to the bottom of a measured length of 12-grain primacord. Once the primacord is dropped in the hole to the desired distance, a second wedge is slid down to create a friction lock with the first wedge at the bottom of the hole. Drill cuttings and small rocks are then dropped on top of the wedge, in order to seal the space around the wedges and create a platform for the explosives. The holes are loaded with 30 pound cartridges of water gel. A primer with a measured delay is assembled, and the primacord is fed through the booster cord-well, which is slid down the hole on top of the explosives. Stemming generally consists of a sand column. Finally, the holes are topped off with water.
Powder factors are ideally less than 1.1lb/ton (0.5kg/t).
There are several stages in the ground support of VCR drifts and stopes. Pillars may be designed in the sill to add stability and minimize widths. Wall stability can also be achieved with secondary reinforcement, the installation of ground support to each blasted round. Long bolts (5 to 8 ft) are installed, often in combination with mats, to immediately secure the back and walls. The most common form of secondary reinforcement is cable bolts, which are used to secure the back and hanging wall during blasting and mucking. Not only does this ensure the safety of the working crews, but it helps to minimize dilution. The length of these bolts is determined by the size of the sill, ground conditions, and future mining plans for ore above the top sill; generally varying from 30 to 60 ft.
Typical VCR cable bolt installation involves twin wire rope cable bolts, installed in parallel and grouted with cement. A standard cable bolt pattern consists of 10 bolts by 10 bolts, with 10ft between rows and bolt points.
Several methods are employed for ground control during blasting, but hanging wall fractures are minimized most effectively by the use of controlled blasting techniques. In addition, by keeping the stope as full as possible with broken rock, the hanging wall is provided with support during blasting.
The backfilling process begins in a stope as soon as the ore within as been completely drawn down. The backfilling process involves the proper coordination of the following factors: availability of waste material, need for mill tailings disposal, crew availability and time. Crews are required to block off scrams and the bottom sill so that backfill material cannot intervene with stoping operations elsewhere in the mine. Walls typically consist of a concrete slab, with a thickness between 18 and 24 inches. Pumps are used to place the concrete in the forms, and stopes are allowed to drain through the installation of “weeping tiles”, which are devices that allow water to pass through, while preventing the passing of sand.
The sand is usually added to the empty stopes to bed the walls and protect the drain lines. Waste rock is then dumped into the stope, unless it is very inaccessible. Ideally, stopes are completely filled with waste rock, and then sand is poured in to fill the voids, however this depends on the amount of waste and time that is available. In the event that there is insufficient waste material and time, sand will be poured into the stope concurrently with the waste rock. If a very competent backfill is required (so that mining can be done against it), concrete is mixed in with the waste material. A period of 6 months is typically scheduled for the water to completely drain out and the fill to solidify. Backfill is placed in the by dumping it through a borehole that leads to the top of the empty stope, or using LHD equipment.
Mining areas that have not been secured by rock bolts pose a high threat to the safety of workers. One reason VCR mining is such a safe method is that workers are only required inside two areas, the top and bottom sills; both of which get rock bolted immediately. There is therefore no need to install ground support immediately after a blast, which saves money and time . An overhand cut and fill operation would require that 13 to 16 lifts be rock bolted, which would be expensive and time consuming. The reduction in required ground support with VCR mining can be up to 88%, which means that workers within a VCR mine are effectively protected at all times from overhead rock falls.
In addition, VCR mines that are well designed have the following safety benefits: good drainage, good visibility, stable ground, smooth corners and good ventilation. As Dr. Walter Curlook, the vice chairman at Inco, says: “improved safety comes hand-in-hand with improved productivity.” 
- Stellman, Jeanne Mager. "Encyclopaedia of Occupational Health and Safety Fourth Edition" Stellman, Jeanne Mager. International Labour Organization, 1998. p. 74.14.
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