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International Journal for Mining, Equipment and Technology – Made in Germany

Glückauf 1-2010

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MINING REPORTER

GLÜCKAUFmining reporterInternational Journal for Mining, Equipment and Technology – Made in Germany

© Copyright: VGE Verlag GmbH, Essen

Publishing House:VGE Verlag GmbHP.O. Box 18 56 20 · 45206 Essen Federal Republic of GermanyInternet www.vge.deE-Mail [email protected] +49 20 54 9 24 - 109Phone +49 20 54 9 24 - 0

Managing Director:Kristian Senn, [email protected] Ullrich-Ley, [email protected]

Chief Editor:Dipl.-Ing. Karsten Gutberlet, [email protected]

Advertising sales:Ute Perkovic, [email protected]

Distribution Manager:Michael Ullrich-Ley, [email protected]

Producer and DTP-Production:Kerstin Finkensiep, [email protected] Stimper, [email protected] Giertz, [email protected]

Printed by:Mönch Medien & Print, Essen, Germany

April 2010ISSN 0176-4101

EDITORIAL

MACHINERYScooptram ST 14 loader is setting new standards for equipment, safety, performance, ergonomics and serviceability in the com-petitive 12 to 16-t market segment Karl-Heinz Wennmohs 6

Implementing the new EU emission regulations: the challenge facing mining equipment manufacturers Günther Robert Saelhoff 11

Transport Technology in Room and Pillar Coal Operations An Overview Marinus J. Pouls 17

30 years of GTA Maschinensysteme – A Lower Rhine success story Karsten Gutberlet 22

SHAFT SINKINGShaft sinking north of the Arctic Circle Oleg Kaledin,Planning, construction and installation Andreas Neff,of the WS 10 mine project for MMC Dietmar Schilling andNorilsk Nickel, Moscow, Russia Andre Marais 26

HEALTH AND SAFETYA Safe Method for Cutting down Production Loss in Case of a Mine Walter HermülheimFire and Andreas Betka 31

PLANNINGHard rock mining the opencast way versus a combination of opencast and deep mining Alexander Hennig, methods Christian Niemann-DeliusAn energy balance sheet and Thorsten Skrypzak 37

METHANEInternational CMM and VAM Project Development Experiences of a Mining Consultant Thomas Imgrund 42

Paul Rheinländer 4


http://www.becorit.de

4 Glückauf mining reporter 1-2010

EDITORIAL

With an exhibition area of over 500,000 m2 bauma has established itself as one of the largest trade fairs in the world. Hardly any

other exhibition in the branch has a wider range of exhibits and no other event is more international. At the last exhibition in 2007 more than half the 3,002 exhibitors came from abroad (49 countries). 155,000 of the more than 500,000 visitors came from abroad (191 countries). The number of exhibitors at bauma 2010 will probably not be less than the above figure. It will probably be difficult to repeat the record number of visitors in 2007. In any case it will be interesting to know the prevailing sentiment among the exhibitors and visitors. Hence bauma is not only a huge market place on an international scale, but also acts as an important sentiment barometer for the branch.

A special feature of bauma is the concept of the partner region implemented for the fourth time this year. Following Latin America, the CIS states and the Arabian peninsula India is the partner country this year. Not much need be said with regard to the potential of India for the branch: with about 440 Mill. t it is the third largest coal producer and with about 1.2 bn in-habitants it has the second highest population in the world, a young population and is the source of hopes for global economic development. Hence an attractive programme is also being prepared for Indian visitors. The guests will include politicians and representatives of the mining industry, who will provide information on development of the mining sector in the course of the Indian special in the afternoon on 21 April 2010 in the bauma forum in the Mining Hall C2. In addi-tion to representatives of the Ministry of Coal the speakers will include managers of the coal companies Coal India, Singareni Collieries and the state-owned ore mining company National Mineral Development Corporation (NMDC).

In addition to India further country specials will deal with Russia, Latin America and South-east Asia. The specials will focus on contributions from the countries to the current situation and future development in the mining industry. The visitors to the events in the bauma forum can thus obtain impressions at first hand with regard to the opportunities and possibilities for business in these regions. The four regions are markets, to which mining equipment from Germany has been sold on an increasing scale in recent years and which exhibit potential for further growth.

India India offers in particular good sales opportunities in the medium and long term. With an annual production of about 440 Mill. t the sub-continent is the third largest coal producer in the world. By 2012 output will have risen to 520 Mill. t. Exports to the country increased more than tenfold by the end of 2008 and will reach a level of about 60 Mill. € in 2009.

RussiaExports to Russia more than doubled to 353 Mill. € be-tween 2005 and 2008. In 2009 a volume of 235 Mill. €

was achieved. Russia will nevertheless remain the most important foreign market.

Latin AmericaLatin America is developing into an emerging region in the branch. Exports to Central and South America increased tenfold between 2005 and 2009 and rose to more than 130 Mill. €. Copper and important mineral raw materials are the main growth promoters.

South-east AsiaThe south-east Asian market depends to a large extent on the Chinese demand for coal and raw materials. However, momentum for growth also emanates from the economies of the other countries in the region. In 2008 Indonesia produced 150 Mill. t coal. The Philip-pines are not yet a coal exporter, but have numerous mineral resources, which will be developed and worked in the next few years.

An important factor in the success of German mining engineering on the world markets is probably the cooperation of the manufacturers with the mining institutes at the German universities.

Three universities highly respected internationally are in the forefront: the oldest mining faculty in the world, the TU Bergakademie Freiberg, the RWTH Aachen established in 1870 and TU Clausthal are collaborating closely in the search for ways to meet the requirements of customers for efficient and safe technology for the extraction and processing of raw materials. These universities will likewise report on their main areas of current research at the week-end in the bauma forum. In addition 14 universities and research institutes from all over the world will describe the latest research results at the “Research live” stand. To indicate aspects of the mining activities along the value creation chain the German Federation of International Mining and Mineral Resources (FAB) will be present at the VDMA stand in Hall C2 at the bauma exhibition.

I hope that as exhibitors and visitors you will have an informative, interesting and successful time at bauma 2010.

Glückauf!

bauma 2010 – branch event, sentiment barometer, market placeDr.-Ing. Paul Rheinländer, Chairman of the Mining Machinery Section in VDMA and managing director of Gebr. Eickhoff Maschinefabrik und Eisengiesserei GmbH, Bochum, Germany

http://www.vdma.org/wps/portal/Home/en/Branchen/M/BBM?WCM_GLOBAL_CONTEXT=/vdma/Home/en/Branchen/M/BBM

Sinking

Drilling and heading

Extraction (underground)

Roof support

Cutting heads

Shearer drums

Conveying and transport

Ventilation and air conditioning

Open-cast mining

Highwall mining equipment

Process control and automation

Central control and stations

Disposal mining

Deep well drilling

Preperation plants

Coke oven and equipment

Up-garding plant

Driving aggregate

Pumps and compressors

Electrical equipment and Control devices

Communication, data processing and transmission

Measuring devices and safety installations

Pneumatic and hydraulic tools

Tools

Safety measures and health protection

Chemical products and build-ing materials

Fittings and valves

Chains, ropes, pipes, wires and wire goods

Rubber and plastic products

Fixing materials, bearing lubrication and other products

Mining companies/contractors

Consulting/Engineering

Machines and plants for process control, process measuring technique

Mining companies

Environmental Protection

Pit gas

Research, theory, technology transfer

Mining special literature, fi lm and photo documentary

Other products

Personal protective equipment

Firedamp protection, fi re prevention, environment protection

General safety at mines

Main exhibit categories

OrganisersOrganisers

11th International Trade Fair for Mining Technology Exploration, Underground Mining, Open-Cast Mining, Mineral Processing and Coal Preparation

7–10 September 2010 Donetsk/Ukraine

Messe Düsseldorf GmbHMesseplatz40474 DüsseldorfGermanyPhone +49/211/45 60-02Fax +49/211/45 60-77 [email protected]

17th International Trade Fair for Mining Technology Exploration, Underground Mining, Open-Cast Mining, Mineral Processing and Coal Preparation

1–4 June 2010 Novokuznetsk/Russia

www.ugol-mining.com

Now with Special Feature Safety & Security

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6 Mining Reporter 1-2010

MACHINERY

Scooptram ST 14 loader is setting new standards for equipment, safety, perform-ance, ergonomics and serviceability in the competitive 12 to 16-t market segment Dipl.-Ing. Karl-Heinz Wennmohs, Atlas Copco MCT GmbH, Essen, Germany

The global demand for high-performance loaders for the deep mining industry, usually known as LHDs (load-haul-dump loader), has undergone

an erratic development over the last ten years in terms of the demands of users and the transport capabilities of the different machine series and manufacturers. The trend towards larger unit sizes has now clearly estab-lished itself, with the 12- to 16-t-class having developed into a lucrative market segment for manufacturers.

Developments in this particular area cannot simply be initiated by the manufacturers alone from a purely marketing perspective, as the end-customer will ulti-mately have to be included in the design process if any project of this kind is to be successfully completed as a builder-user partnership.

Looking back

When the mining industry talks about mobile load-ing and transport machines we generally hear names like

LHDs, ñLoaders, ñMuckers, ñBoggers, ñScoopy, ñThe Wagner and ñThe Scoop. ñ

But there is one name that has now established itself worldwide – and that is “Scooptram”.

Developments in the raw materials market have dramatically driven up demand for underground loaders in recent years. At the same time mines have opened up greater opportunities for LHDs in the larger size range. The 12- to 16-t-class is now an impor-tant sector of the global LHD market and Atlas Copco has developed a completely new machine for this particular segment. The decisive criterion for LHDs is perform-ance in tonnes per shift. If identical power categories are being matched one against the machine with the higher productivity will soon win out. As well as pure perform-ance, factors such as operating costs and machine availability are also important criteria in the selection process. All these features, plus the ergonomic constraints imposed by the operator’s workplace, were taken into account during a collaborative design process involving a number of global mining companies – and the result was the Scooptram ST 14.

During the period 1940 to 1955 the Elmer brothers and Eddie Wagner developed diesel-powered loading and transport vehicles for surface worksites.

During this time experience was built up in the design of four-wheel drive and articulated steering

Figure 1. The first underground

LHD – the Scoop MS-1.

http://www.atlascopco.de/dede/

7Mining Reporter 1-2010

MACHINERY

systems. In 1958 Eddie Wagner founded the Wagner Mining Scoop Company and presented his first mobile loader for the deep mining industry. Designated the “Scoop MS-1” this machine was the forerunner of today’s generation of LHDs (Figure 1).

The new Scooptram ST-5, which went into service in 1963, heralded the market launch of the LHD concept. This range of machines led to the successful worldwide introduction of LHD technology during the years 1963 to 1975 (Figure 2).

The Wagner company took out a large number of technical patents over the years, including the design for the E-o-D (Eject or Dump) bucket. The SAHR (Spring Applied, Hydraulically Released) braking system was also a key development from this period.

Wagner was taken over by Atlas Copco in 1989. In 2002 the firm relocated from Portland, Oregon, USA, to Örebro in Sweden and was renamed Atlas Copco Wagner.

Design and development

A global market and demand analysis was first car-ried out for the development of the new Scooptram ST 14 range. This identified a clear trend towards a shift in LHD sizes in the underground mining industry. Machines in the 6- to 10-t-category were being sup-planted by equipment in the 10 to 15-t class, while turnover in the 18- to 22-t-category was stagnating or even switching to smaller sized machines. In the case of the smaller types of machine this can naturally be attributed to the fact that larger and more powerful loaders are being introduced in order to increase the output per working unit, although the trend that has

been observed in the heavyweight class is harder to explain. The reasons for it lie partly in the operating conditions, while in some cases it has been found that the vehicle components, powertrain and tyres have reached the limits of their operating capacity.

The technical capabilities of an LHD vehicle are primarily measured by its loading and transport ca-pacity per shift. A loader whose haulage potential is

Figure 2. Scooptram ST-5.

Figure 3. Cut and fill stoping.

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MACHINERY

After numerous field trials at various mines the ST 14 was declared ready for series production and was then purchased by mine operators around the world for use in typical LHD applications, e.g. for cut and fill stoping (Figure 3), loading ore and bringing in back fill material, as well as for sub-level stoping and room and pillar mining methods.

All these typical mining applications assume a defined maximum distance for the LHD transport cycle. What this distance will be will depend on a number of factors and will naturally differ from one mine to the next.

Machine performance

Getting the optimum performance from an LHD vehicle essentially depends on how quickly and readily it can fill the bucket. This operation is very aptly described as “one-pass loading”, which simply means that the loader with its power and optimal scoop design penetrates the muck easier in a single pass, loads the bucket and then continues on to the dumping point (Figure 4).

Another important factor for optimum loader de-ployment is operator visibility in the loading direction and when reversing at speed.

A cab height of 2,550 mm affords the driver an excellent line of sight in both travel directions over the top of the engine compartment and bucket, which are 1,980 mm in height (Figure 5).

The load sensing system fitted to the hydraulic cir-cuit provides about 44 % more available performance compared with similarly powered machines with open centre hydraulic systems, which is particularly useful when digging. This system can deliver a 14 % increase in travel speed on ramps with a 15 % gradient. This increased speed, combined with faster filling of the bucket, are the factors that make for a higher loading performance.

With the load-sensing system the hydraulic flow is determined by the actual demand and ”oil leakages” are reduced to minor adjustment flows. This reduces the energy input and the motor consumes about 10 % less fuel than similarly powered units with open centre hydraulic systems.

About 20 years ago the first Atlas Copco under-ground drilling machines were fitted with a CAN bus system for control purposes. Over the years this technology also came to be used on other product ranges. It was therefore decided at the planning and design phase that the CAN bus control system (or RCS – Rig Control System) should also be fitted to this new generation of loaders.

The hardware and software installed on the Scoop-tram ST 14 provides an ideal platform for ongoing machine automation.

Safety and ergonomics

As part of the pilot study a new generation of opera-tor compartments was built specifically for this new model range. These ROPS and FOPS tested cabs feature a larger window area for improved operator visibility and noise levels are now reduced to a maximum of

Figure 4. Bucket design.

technically smaller than that of a larger machine can however prove to be a better solution if it is faster and more mobile in its movements, with the result that it can deliver a significantly higher shift performance.

And this also applies when viewed from the other perspective: the much larger vehicle with its relatively slow movements will deliver a lower output per shift and will therefore be inferior to the smaller machine.

In addition to these fundamental considerations factors such as diesel consumption and ergonomic/operator-friendly design are now coming increasingly to the fore. This is covered in a separate section.

The demands of the market, a comparative review of the different types of loader currently in service and the need for a real improvement in the existing systems and technical features led to the development of the 14-t Scooptram ST 14.

Figure 5. Scooptram ST 14 – technical data and dimensions.


MACHINERY

80 dB(A). The operator is now exposed to much lower levels of vibration, a factor that was greatly appreciated by loader drivers during underground field trials, along with the improvement in machine performance.

If the cab door is not fully closed the braking systems are switched to parking mode, the steering cannot be

activated and the bucket and boom are also hydrauli-cally locked (Figure 6).

The Atlas Copco foot box is another significant comfort factor for the operator during his long stints on board the vehicle (Figure 7). The new box fea-tures hanging control pedals that ensure maximum

Figure 6. The loader cannot be operated with the cab door open.

Figure 7. Atlas Copco foot box.

Figure 8. Fields of vision from the new operator’s compartment.

Please visit us at bauma 2010, in hall C2, booth 308 and hall C3, booth 311

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MACHINERY

legroom and allow the driver to stretch out his legs for maximum comfort.

Visibility from the operator’s compartment was greatly improved as a result of simulation exercises carried out during the Scooptram ST 14 planning and design phase (Figure 8).

The benefits of the new cab design were very positively received by LHD drivers.

Servicing and maintenance

Servicing and maintenance experience acquired from mines around the world was taken into consideration when planning the ST 14 loader and this is reflected in the design of the new vehicle.

All key maintenance routines, such as filter chang-ing, valve inspections and oil level checks, can now be undertaken ”on the ground”.

The RCS control system is able to identify the re-quired maintenance work and log the servicing data for future reference. It also provides extensive diag-nostic support for troubleshooting with information displayed on the on-board display screen.

This information can additionally be transferred to the mine’s existing data system and in this way can be used for scheduling maintenance work while at the same time performing fault analysis and trouble-shooting routines.

Outlook

Fitted with the latest generation of diesel engines to comply with current emission regulations the new se-ries of LHDs in the 12- to 16-t-class has been designed to meet the demands of today’s mining industry.

Everyone involved in the development of the new range is fully aware of the vehicle emission limits that are to be imposed on underground loaders and dumpers in the run-up to 2015. A typical example taken from the dumper sector illustrates just how tough these targets are:

Current vehicle emission limits in 1995 for single dumper vehicles will be equivalent to the emissions from 85 similarly sized dumpers from 2015 on.

This fact represents a real challenge for motor and vehicle manufacturers and it is one that will have to be solved within a relatively short period of time. With its forward-looking design the ST 14 loader is already geared up for future emission reduction targets.

LHD operations typically involve repetitive cycles over the working shift. At the extraction face this means taking material from the muck pile and transferring it to a continuous conveyor or drop-hole, while in stope filling operations it will involve collecting the back fill material (from the drop-hole or bunker) and transport-ing it along a roadway to the discharge point.

In critical areas the loader can also be used in remote-control mode so that the driver can operate the machine by radio from a safe distance. Advanced on-board video camera technology is now available that allows the loader to be remotely operated from practically any part of the mine.

But the real breakthrough involves the progressive automation of the loading and transport cycles. The first step in this direction will see the machine making the trip from the loading point to the tipping area automatically and without any assistance. The actual loading and dumping cycle will be completed manu-ally by remote control. Such operations could also be continued during shift changeovers when blasting is carried out.

The key factor in all this is the need for an on-board operator and everyone involved in developing this new technology is working towards the ultimate objective of the fully manless LHD cycle.

The first steps have now been taken in this direction with the Scooptram ST 14, whose design remit included this development objective from day one.

The automation system should be hands-on, tailored to the task and highly functional. All the hardware components required to operate the au-tomatic control system are now on the machine. All that is needed is the availability of and connection to the mine’s own performance-capable Wireless LAN network (Figure 9).

Excellent results have already been achieved with the Scooptram ST 14 set up to deliver back filling mate-rial, in other words permanently assigned to carrying out a repetitive working cycle.

Further deployments in various mines around the world have confirmed this. While a number of require-ments have yet to be met – for example there is still a catalogue of questions surrounding the actual ”one pass loading system” – it seems likely that the experi-ence acquired to date will be sufficient to resolve any outstanding problems in this area.

Figure 9. Scooptram ST 14 equipped for automatic operation.


MACHINERY

Implementing the new EU emission regulations: the challenge facing mining equipment manufacturersDipl.-Ing. Günther Robert Saelhoff, GHH Fahrzeuge GmbH, Gelsenkirchen, Germany

Controls on emissions from diesel vehicles are now being applied in the off-road sector with the introduction of the next stage of the emis-

sion reduction programme in the EU, in North America and in Japan. The switchover from EU Stage III A to EU Stage III B, and from Tier 3 to Tier 4 Interim, will have a much greater impact on motor and engine builders and original equipment manufacturers than the gradual introduction of the previous emission stages.

For engines with a power rating above 130 kW Stage III B is due to come into force at the turn of 2010/2011, while for the 56 to 130 kW range the new regulations will take effect in late 2011/early 2012. The permitted limits for:

Vehicles for today’s deep mining industry are now subject to the same exhaust gas restric-tions as their off-road counterparts above ground. The space constraints imposed on mining vehicles mean that considerable ef-fort and expense is needed to fit the coolers and other components needed for the ex-haust-gas aftertreatment system. By taking large-capacity LHD vehicles as an example the paper examines the kind of constructive measures now being contemplated so that even mining machinery of this kind can be equipped with SCR technology.

NO ñ X (nitrous oxide), CO (carbon monoxide), ñHC (hydrocarbons) and ñPM (particulates) ñ

are shown in Figure 1. There are no limit values for CO2 (carbon dioxide) (1).

While the limits imposed by Stage EU III A/Tier 3 Interim could be achieved by engine-internal measures, EU III B and Tier 4 Interim will require a combina-tion of engine modifications and additional exhaust gas aftertreatment. The available engine-internal measures and options for exhaust gas treatment are depicted in Figure 2.

There are as yet no tried and tested technologies available for underground vehicles that are capable of meeting the threshold limits imposed under EU Stage IV and Tier 4 Final.

All engine-internal measures result in more efficient motors with a smaller combustion chamber, which in turn means an increased cooling requirement.

Supercharged engines with intercooling not only give reduced NOX emissions but also have a longer operating life, while there is a significant increase in torque at lower engine speeds. This makes for a higher power density. For motor manufacturers this means that more cooling capacity is required and en-gine braking is much reduced because of the smaller combustion chamber.

Exhaust gas recycling is a technique frequently used by engine builders in order to reduce nitrous

Figure 1. Permitted limits for NOX, CO, HC and PM.

http://www.ghh-fahrzeuge.de/index.php?lang=english


MACHINERY

either using engine-internal NOX reduction in the form of cooled exhaust recycling in combination with a diesel particulate filter or fitting an SCR system (selective catalytic reduction). In this case the diesel particulate filter (DPF) is designed to reduce soot particle levels.

Regenerating the diesel particulate filter requires an electronically controlled fuel injection system that can carry out a series of injections per working cycle at very high injection pressures. The burner itself takes up a certain amount of space and has to be fitted upstream of the DPF right at the intake to the exhaust collector pipe. Post-engine NOX reduction using SCR technology requires a urea solution as an NOX reduc-tion agent and according to current practice this has to be fed in using a dosing system. The urea solution

oxide levels. The best results are generally obtained by employing a system known as ”cooled exhaust gas recycling” (Table 1). This process also requires additional cooling capacity (3).

As well as providing a larger cooling surface ad-ditional space also has to be found to accommodate the exhaust gas aftertreatment components. This means a higher exhaust emission stage will call for a larger number of components and a greater amount of space in order to achieve the legally required result. The problem of providing sufficient space for cooling and exhaust gas treatment is particularly critical for underground machinery, which usually has to oper-ate in confined conditions. Meeting the requirements imposed by EU Stage III B and Tier 4 Interim means

Figure 2. Available engine-

internal measures and options for

exhaust gas treatment.


MACHINERY

known as Ad Blue is commonly used on roadgoing vehicles. This liquid is injected into the exhaust line upstream of the SCR catalyst and a hydrolysis reaction releases ammonia and CO2 from the urea solution. The ammonia can react with the nitrous oxides in the exhaust gas when the temperatures in the SCR catalyst

reach a certain level. Urea consumption is between 2 and 8 % of the amount of diesel fuel used. The vehicle must therefore have an on-board tank of sufficient capacity (4) (Figure 3).

The feed tank for the urea solution must be sited in a well protected area and the urea solution has to

Engine cooling Cooling water Charge air

Vehicle/Motor Output

[kW/rpm]

Vol.-flow

[L/min]

Intaketemp.

[°C]

Coolingoutput

[kW]

Max. press.loss

[mbar]

Max. op.

press.[mbar]

Mass.-flow

[L/min]

Intaketemp.

[°C]

Out.Temp.

[°C]

Coolingoutput

[kW]

Max. press.loss

[mbar]

Max. op.

press.[mbar]

SLP12(COMIII)/ TCD2015VO6

240/2,000 410 103 125.0 317 1.0 0.520 172 50 64.7 70 3.0

SLP12(TierIVi)/TCD2015VO6

240/2,000 410 103 108.0 317 1.0 0.540 157 50 59.0 70 3.0

SLP12(TierIVi)/ TCD2013VO6 4V

240/2,000 314 110 153.6 171 1.0 0.267 184 46 37.6 100 2.5

LF-7.4(COMIII)/ TCD2013VO6 2V

181/2,300 198 110 90.6 230 1.0 0.301 189 50 42.7 100 2.5

LF-7.4(TierIVi)/ TCD2013VO6 4V

180/2,200 363 110 140.4 229 1.0 0.252 171 46 32.1 100 2.5

LF-7.4(TierIVi)/ TCD2013VO6 4V

180/1,900 314 110 124.2 171 1.0 0.238 164 46 28.6 100 2.5

Table. 1. Motor cooling – as the exhaust gas levels are reduced the cooling requirement of the engine will increase as a function of the exhaust treatment process. A higher cooling capacity can only be achieved by using larger-sized units for the water cooling and intercooling systems.

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MACHINERY

be kept at a temperature of between –5 and +25 °C. When the vehicle is operating this temperature range must not exceed –11 to +35 °C in order to prevent crystallisation. Even this relatively small range poses problems at many mines and additional cooling may be required for the urea tank.

The urea solution has to be metered out at a rate that is generally equivalent to the NOX emissions. If too much urea is fed-in the ammonia formed from it can-not react with the NOX and so will be released into the environment. This creates an extremely pungent smell, especially in enclosed areas, and may also be damaging to health. This problem can be resolved by incorporat-ing an additional oxidation catalyst – the so-called CU (clean-up) catalyst – behind the SCR catalyst.

The individual exhaust gas aftertreatment systems are shown in Figure 4. The burner required for the re-generation of the diesel particulate filter is connected into the exhaust flow upstream of the filter, as stated above. Vehicles that have to operate on steep inclines are often fitted with an ”engine retarder”.

The combined use of burner and DPF means that a number of motor manufacturers are no longer able to include an engine retarder in the exhaust gas line. In such cases the companies concerned have to fit a differ-ent version of the auxiliary brake retarder, which also requires additional space. Dispensing with the engine retarder would again call for extra cooling capacity to dissipate the braking energy. An effective solution here is to store the braking energy and to use it when increased output is required. This can for example be achieved by pressure charging hydraulic accumulators by way of a volumetric flow from the duty pumps, provided these have the requisite control system. Approximately 0.08 kWh of energy can be stored per braking cycle over a period of about 8 s. Assuming 30

cycles per hour and around 200 working days a year this means an energy saving of around 7,000 kWh or 2 – 3 % of the total energy requirement.

A fullscale SCR exhaust aftertreatment system will take up several times more space than a conventional exhaust gas silencer. The following components will be required, depending on the particular emission stage or threshold in question:

Components in the exhaust gas flow whose posi- ñtion is fixed because they must be located either in the hot zone of the exhaust system (DOC, DPF) or in the cooler zone after the urea injection stage (SCR, CU CAT). The distance between the individual components is also effectively predefined by the fact that certain values have to be met on the engine test bench. Depending on vehicle type and size the need to comply with EU regulations on exhaust emissions could mean that it would no longer be possible – at least from an economic viewpoint – to build certain models.

Compared with a conventional silencer unit an exhaust aftertreatment system requires a significant outlay on engineering, and a lot of space, in order to achieve the emission limits imposed by EU stage III B/Tier 4 Interim. Figure 5 shows a typical exhaust aftertreat-ment system with particulate filters and SCR catalysts, as fitted to a Deutz diesel engine type TCD 1015 V08. As can be seen, there is practically no access to the engine itself and servicing components like the oil filter have to be relocated elsewhere, which in turn increases the maintenance costs.

The components listed below are required in order to meet emission limits laid down in Stage III B/Tier 4 Interim:

Cooling: ñAdditional engine cooling. í

Figure 3. Vehicle fitted with

2015 Com III A motor and

vehicle fitted with 2015 Tier IV

motor.


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Intercooling. íCooled exhaust gas recycling. íAuxiliary brake retarder: ñRetarder, hydraulic hybrid or similar. íExhaust gas aftertreatment: ñBurner for particulate filter regeneration. íDiesel oxidation catalyst. íDiesel particulate filter. íSCR catalyst/clean-up catalyst. íAd Blue (UREA) tank. íInjection unit. íMonitoring/control electronics, signalling system. í

One way of reducing the cooling requirement and the size of the exhaust system is to “downsize” the diesel drive motor.

In this case the power output, and ideally the di-mensions of the engine too, can be reduced without incurring any loss of performance from the actual machine. This calls for targeted modifications to the different systems fitted to the vehicle. While each individual measure only produces a relatively small impact, the sum total of such modifications can result in a significant saving in power demand. By carrying out an efficiency analysis of the individual

Figure 4. Exhaust gas aftertreatment systems.


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systems it is easily possible to achieve savings of well above 10 %.

Drive systemOverall LHD efficiency can for example be improved by opting for a hydrostatic drive instead of the stand-ard hydrodynamic drive system. In such cases quite significant energy savings can be achieved on each loading cycle and this is reflected in a much lower cooling requirement and significantly improved fuel consumption.

Fewer pumps required for the drive hydraulics

By priority switching or series switching the different load consumers it is possible to reduce the number of pumps units and hence improve the overall ef-ficiency.

Variable displacement pumps:By employing demand-driven variable displacement pumps in the steering and working hydraulics efficiency levels can be improved during the part-load and no-load phases, as on a hydrostatic drive system.

Braking energyOn hydrostatic drive systems, for example, it is possible to store the braking energy by building-up pressure in a series of bladder accumulators and then using this to drive the vehicle when it is in travel mode. While

the amount of energy that can be stored from each braking action is relatively low, significant savings can be obtained from the sum total of all the LHD load-ing and travel cycles. However, the most important benefit of the hybrid braking system is that it relieves the strain on the friction brake and helps extend its service life.

Auxiliary drivesBy adjusting the controlled fan drives to meet the current cooling requirement it is possible to achieve a substantial reduction in the fan output. Fitting insu-lated driver’s cabs, ensuring good temperature man-agement and reducing the heat input into the driver’s compartment are all measures that can significantly cut the output of the air-conditioning system.

If a smaller-sized engine can also be fitted this not only makes for a reduced engine volume but also allows a more compact air filter unit to be installed. A lower volumetric flow in the exhaust gas line also means that in certain cases a much smaller exhaust system and aftertreatment unit can be fitted. The reduced demand from the cooler makes for smaller-sized cooling units, a lower fan performance and smaller fan drive components. Finally, the reduction in fuel consumption means that the vehicle can operate with a smaller diesel tank (Table 2).

Operating experience with the Deutz TCD 1015 V08 diesel engine in a 20-t LHD fitted with an exhaust gas aftertreatment system has shown that exhaust aftertreatment using a combination of particle filter and downstream SCR catalyst with Ad Blue injection on the vehicle can only be achieved under consider-able space constraints. Compromise solutions have already adopted in terms of the accessibility and serviceability of certain components. On smaller ma-chines and in more confined operating conditions it would not always be possible to install an elaborate after-treatment system.

Underground trials at the Zielitz mine (K + S Kali GmbH, Germany) with an 18-t LHD type LF-18 fitted with a TCD 1015 V08 engine and a Johnson Matthey SCRT system were generally successful and the compo-nents sited along the exhaust gas line stood the test well. Only the compressor and metering system were found to require improvement. Another machine from the same vehicle range was subsequently fitted out from new with the same SCRT system and has now been in regular service for a number of months.

Operating experience indicates that, at least as far as mining machines of this size are concerned, the limits imposed by EU Stage III B/Tier 4 Interim are both achievable and practicable when using SCRT exhaust aftertreatment systems installed downstream of the EU Stage II engine. This will have to be verified by long-term tests.

References1. Nonroad Diesel Engines, Emission Standards, Dieselnet.2. Deutz AG: Variable Emissions-Reduktions-Technologie.3. John Deere: Emissionstechnologien.4. Eidgenössisches Department für Umwelt, Verkehr und Energie: Emissionsvorschriften Schweiz und International.

Figure 5. Typical exhaust aftertreat-ment system with particulate filters and SCR catalysts as fitted to a Deutz diesel engine type TCD 1015 V08.

Table 2. What are the benefits of reducing the motor power output?

May be possible to use a smaller model of engine

Engine size is reduced

Reduced air requirement Smaller air filter unit

Lower exhaust gas flow Smaller exhaust system

Lower cooler requirement Smaller cooler

Reduced fan output Smaller drive components

Lower fuel consumption Smaller fuel tank

In spite of the reduction in engine power the systematic use of measures to reduce power loss will result in a vehicle of comparable performance.


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Transport Technology in Room and Pillar Coal Operations An OverviewDipl.-Ing. Marinus J. Pouls, Bucyrus Europe GmbH, Lünen, Germany

One of the lesser known methods for mining coal in Europe is Room and Pillar (also called Bord and Pillar). This method is extensively

used in South Africa (90 % of UG production), India (95 %) and USA (50 %).

A room and pillar section consists of roadways and cross-cuts between the roadways. A typical panel layout is shown in Figure 1.

The pillar sizes between the roadways and cross-cuts vary depending on mining height, roadway width, seam depth and whether depillaring will be under-taken or not. For example in a mine having a depth of 150 m, a mining height of 3.5 m and a road width of 6 m the pillar size would be 15 m x 15 m, in case no depillaring takes place or 20 m x 20 m if depillaring is planned. Mining can take place in seam heights as low as 0.9 m (Figure 2 for low profile Continuous Miner) and up to 5.5 m. Road width varies between 5 and 6.5 m. Room and pillar operations prefer flat condi-tions or seams with only a slight gradient (preferably less than 1 in 6 or 9°).

This method consists of the following activities:Cutting of coal and loading it on to a transport ñvehicle with a Continuous Miner.Transporting the coal to a feeder breaker with ña coal hauler (Figure 3).Crushing and loading coal on to a belt conveyor ñwith a Feeder Breaker (Figure 4). This consists of a hopper, a chain conveyor to transport the coal through a crusher onto a belt (capacity up to 1,200 t/h).

The present paper will give an overview of the actual status of the coal hauling in the underground room and pillar mines. It will give a general overview of room and pillar operations and will describe in more detail the different coal hauling alterna-tives available, including the advantages/disadvantages of each system. The coal haulingalternatives are:

Batch transport with Shuttle Cars, ñDiesel Haulers and Battery Haulers ñContinuous transport with Continuous ñHaulage Systems

It will further describe operational aspects of coal hauling in operations, such as cable layout for Shuttle Cars. It will furthermore highlight improvements made in the past years and those which will or are being imple-mented at present or in the near future.

Supporting roof and if required ribs with a ñRoof Bolter (Figure 5). The Roof Bolter can have one, two or four bolting rigs. It further has a temporary roof support to ensure a safe working environment.

Figure 1. Typical Panel Layout.

Figure 2. Continuous Miner in a thin seam operation and discharge end.

Figure 3. Shuttle Car.

http://www.bucyrus.com/mining-equipment/room-pillar.aspx


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Coal Haulage

The interesting aspect is that the principals of coal haulage in a room and pillar section have not changed in the past 50 to 70 years. The first shuttle cars were introduced already in 1938 by Joy and battery haul-ers as early as 1950’s. The improvements have been mainly in availability and quality.

The main aspect of this presentation is on the coal haulage between the Continuous Miner and the Feeder Breaker. This can be in two forms:

Batch transport: ñShuttle Cars (Capacity up to 30 t). íCoal Haulers (Capacity up to 22 t). íContinuous Transport: ñWith belt conveyors. íWith chain conveyors. í

Batch transportsBasically, there are two versions, each with different alternatives

Shuttle CarThe Shuttle Car as the name indicates “shuttles” be-tween the Continuous Miner and the Feeder Breaker. It is the most used transport system for hauling coal. According to the information available more than 2,600 units are in use or more than 70 % of the total population of coal haulers.

There are two models:Electrical powered with a cable. ñElectrical powered by having a battery on board ñ(only used in relative small numbers).

The main characteristics for the cable powered version are (Figure7):

Coal is loaded on one side and the internal chain ñconveyor moves it forward to the front.Requires a cable which is spooled onto a cable ñreel for it’s energy supply. Location of the cable reel is on the LH- or RH- ñside. The driver is located on the other side. When operating SC’s the cable routing in the section plays an important role. The routes of the SC’s must be independent from each other (Figure 8).The driver can seat perpendicular (straddle) to ñthe direction or in line with the direction (two seat option).Can be equipped with a discharge boom or ñnot.Does not have to turn when travelling between ñContinuous Miner and Feeder Breaker.Is reported to have the cheapest operating cost ñ(75 to 80 % of a battery hauler).

Coal HaulersThe coal haulers can be identified by the fact that they consist of two parts with articulation between the two, the parts are:

Trailer for transporting the material ñTractor with the “power” unit ñ

Coal Haulers can be powered by (Figure 9):Battery (over 300 in use). The standard process ñis to have two spare batteries: whereas one is

Utility activities such as cleaning the floor and ñtransporting material (Figure 6). There are in principal two versions: Battery powered “scoop”. íDiesel powered utility vehicle (similar to a LHD). í

Figure 4. Feeder Breaker.

Figure 5. Twin boom Roof Bolter.

Figure 6. Scoop and Utility Vehicle.

Figure 7. Shuttle Car showing

cable a position of driver.


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being charged the second battery cools down. Under normal operating conditions the batter-ies last for 8 h.Diesel engine (over 100 in use) ñ

Some general characteristics:Has to turn when going to the Feeder Breaker ñto be able to discharge.Is heavier than the SC but has a large volume ñfor transporting coal.It has a push blade to discharge the material ñ(Figure10).The driver sits on the side of the machine ñ(Figure 11).It is possible to operate more than two machines in ña section. The machines can drive in a circle. With a Shuttle Car this is not possible.

Continuous transportAs the name states the Continuous Miner does not have to stop cutting because the material is hauled continuously away to the Belt Conveyor. There are different types of haulage systems:

The modular designed continuous haulage ñsystem (Figure 12).The single strand continuous haulage system ñ(Figure 13).

The Modular designed continuous system consists of short individual “haulage” systems combined into one. The single strand system has one through going “haulage” system.

Figure 8. Cable layout in panel.

The continuous haulage system as shown in Figure 12 has following components:

Mobile Bridge Carrier. ñPiggybacks, also called Mobile Bridge Con- ñveyors.

Every Mobile Bridge Carrier has a driver. Some general characteristics:

Is more complicated to operate. ñAdvisable to have two Roof Bolters in the sec- ñtion, one on the LHS and one on the RHS.The panel layout of a Continuous Haulage ñsystem is slightly different from a “standard

http://www.voithturbo.com/startup-components.htm


MACHINERY

Product comparison

When choosing one of the haulage systems the first most important question is which of the above men-tioned haulage methods is the most applicable one. As usual there is no black and white statement pos-sible. A number of aspects play a role in determining, which method is the most favorable one considering the available operating conditions. These are but not limited to:

Geological aspects (type of floor, mining thick- ñness, inclination).Mining/operational aspects (pillar sizes/traveling ñdistance, flexibility, investment requirement, operational cost, ventilation).Local experience. ñ

Hereafter some of these will be discussed.

GeologyFollowing aspects have an influence on the decision making process:

Inclination. ñIn steeper conditions (more than 8° or 1 in 7) the íDiesel Hauler has an advantage over the Shuttle Car and Battery Hauler. Battery haulers are preferably only used in relatively flat conditions because bat-tery life reduces when they have to work in inclined conditions. Continuous Haulage is more difficult to operate under inclined conditions more than 6°.Type of floor. ñIn soft floor conditions the Shuttle Car and Diesel íHaulers are preferred. The Continuous Haulage requires a relative good to semi good floor.Mining thickness. ñIn low seam applications (i.e. less than 1.5 m) bat- ítery haulers have an advantage over Shuttle Cars. In seam heights less than 1.2 m the capacity of the coal haulers becomes low and the use of continu-ous Haulage systems has a positive effect on the haulage/production capacity.

Mining/operational aspectsThe following mining/operational aspects have an influence:

Pillar sizes/tramming distances. ñThe size of the pillars will depend on the depth of the íseam, road width and the mining height. Therefore at greater depth the pillar sizes become bigger and automatically also the tramming distances. Shuttle Car tramming distance depends on the maximum capacity of the cable reel (this is related to cable type used) standard capacities are in the range of 150 to 200 m, the maximum is around 250 m. Con-tinuous Haulage systems have a maximum length of around 120 m. If longer tramming distances are required the battery or diesel powered haulers are the preferred solution.Flexibility. ñShuttle Car and Continuous Haulage systems are íless flexible than the battery powered and diesel powered haulage systems.Investment cost. ñThe cheapest solution is the use of Shuttle Cars, ífollowed by the battery haulers (they require two

Figure 9. Diesel and battery Haulers.

Figure 10. Push blade for

discharging coal.

Figure 11. Location of Driver.

Figure 12. Modular Continuous Haulage System.

section”, the angles of the cross cuts are set at 60° to facilitate a better handling of the system (Figure 14).


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additional spare batteries and a charging station). The Diesel Hauler investment cost is comparable to the battery hauler. The Continuous Haulage system costs the most.Operation cost. ñIt is reckoned that the SC is the cheapest to operate, ífollowed by the battery hauler and then the diesel hauler. The operational cost of the Continuous Haulage system is said to be between the Shuttle Car and Battery Haulers. However, it should be mentioned that the haulage cost do play a major cost but it is estimated that it still only count for about 15 to 20 % of the total operational cost in a section.Ventilation/environment. ñIn case of using diesel powered vehicle special at- ítention has to be paid when using them. Besides the discharge of harmful gasses such as CO, CO

2 etc. also the additional heat generated by the diesel engine in the headings must be considered. But having said this, the effect is not as dramatic as maybe believed.

Local experienceAn important role in the decision which equipment will be chosen plays the local experience at the mine and or adjacent mines. It is not uncommon that a system operates very well in an area but not at all in another. It has been observed that in a mine in one section the continuous Haulage was a success but in the second section it only had problems. The difference was that the first section had a crew with more than 10 years of experience with the system and was also convinced that it was the best they have.

Future and present developments

It is outside the scope of this paper to give a detailed overview. We therefore present you a short enumera-tion of the present and future developments:

Introduction of computerized controls (displays) ñ(Figure 15).Introduction of VFD (Variable Frequency Drive) ñdrive systems.AC drive on Battery Haulers. ñUse of joy sticks for steering the equipment. ñImproved diesel engines ñIntroduction of automatic controlled Continu- ñous Haulage systems

Conclusions

Coal Haulage systems play a very important part in a Room and Pillar section. There are a number of alterna-tives available to an end used. The determination of the most appropriate system depends on the mine.

References1. N.N.: Section by section/Equipment – Underground Coal Haulage. Mining Magazine, September, 2007.2. Stefanko, R.: Coal Mining Technology Theory and Practice.1983.3. Hargraves, A.J. ; Martin, C.H.: Bord and Pillar methods: in Australasian Coal Mining Practice. 1993.

Figure 13. Single Strand Continuous

Haulage System.

Figure 14. Panel layout showing cross cut angles at 60°.

Figure 15. Compartment showing Joy stick control and display.

4. Sanda, A.P. : Haulage 1998 (Underground). Coal Age, September, 1998.5. Sanda, A.P.: Continuous Haulage a Bridesmaid Still. Coal Age, 1989.6. Torre, D.C.: Development in Continuous Haulage for Coal Mining. Mining Engineering, 1982.7. The Battery Powered option. World Coal, 1998.8. Bucyrus Internal Documents on Coal Haulage.


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30 years of GTA Maschinensysteme – A Lower Rhine success storyDipl.-Ing. Karsten Gutberlet, Chief Editor, VGE Verlag GmbH – Verlag Glückauf, Essen, Germany

If you want to hear or write something positive in the current financial and economic crisis then you need to look to the SME sector as the most robust

part of the national economy, for over the years it has been the adaptability of small and medium-sized companies to the fluctuating fortunes of the economy that has helped secure the comparatively high stand-ard of living that we all enjoy. SMEs have been a “tower of strength” amidst the turbulent events that have surged through our economy and the German word “Mittelstand” – or “small and medium-sized enterprises” – has now taken its rightful place amidst the Anglo-Saxon terms that litter the vocabulary of international economists.

SMEs and SME productivity is a subject that is usually spoken about in general terms and rarely do we hear about individual companies or see the faces that run them. Important innovations that emanate from relatively small businesses are often overlooked or take many years to be recognised or even accepted by the established marketplace.

One such example from the German Lower Rhine area is Hamminkeln-based GTA Maschinensysteme GmbH (GTA), a company that owes its origins to the mining industry and currently manufactures equip-ment for the “upper tier” in roadway drivage and tunnelling operations.

GTA was founded by Helmut Heisterkamp in 1979. As a salaried engineer he had become familiar with the confined operating conditions of the German coal industry and he therefore set out to develop and manufacture his own equipment in order to improve the lot of those working below ground (Figure 1).

Mine requirements, transport systems and support technology – the German abbre-viation for this is GTA – were the business activities of the company established in 1979 by Dipl.-Ing. Helmut Heisterkamp, the begin-nings and success of which are closely linked to the coal mining industry. Nowadays the company concentrates on road drivage and transport technology for the international mining and tunnel construction industry.

Suspended machine systems for roadway drivages

The company’s technical and economic success con-tinues to be based around suspended machinery and equipment for roadheading operations in the coal industry. These systems make use of the upper levels in these confined working areas and have now become a key requirement for ensuring productivity and workplace safety. They have also significantly reduced the cost of the preparatory work needed to develop a new coal face.

The introduction of this technology also led to a significant reduction in accidents in this area of opera-tions. The machines and components are continuously being refined and improved and the “GTA platform” has for years now always been pencilled-in as standard equipment when planning new mine drivages. In the major mining markets of China and Eastern Europe, on the other hand, this particular technology initially failed to make any inroads because of the low safety priority and cheap labour costs prevailing in those countries. However this situation is now fortunately changing. The main products that are still being used by the coal industries in Germany and elsewhere are:

Mobile drilling and working platforms. ñRoadway support machines and arch erec- ñ

tors.Drilling and rockbolting systems for high level ñ

work in mine roadways.Roadway support manipulators. ñRoadway start-up platforms. ñRockbolting platforms and slewable working ñ

platforms.The company could have rested on the success that sustained it – and continues to reap rewards even today – over the course of the slow decline in the German coal mining industry. This was the approach adopted by some suppliers that went through a similar

Figure 1. Dipl.-Ing. Helmut Heisterkamp, partner and managing director of GTA Maschinen-systeme GmbH.

http://www.gta-maschinensysteme.de/en/index.asp?p=116

http://www.vge.de


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period of prosperity – and they are no longer around to tell the tale.

GTA always realised, however, that while the Ger-man coal industry provided an outstanding basis on which to build a business, this was a market with a limited shelf life. The company management in particu-lar knew that the principle of dividing the mechanised roadheading equipment into two operating levels, with the economic and safety benefits that this would bring, would also be of real interest to the tunnelling sector. Efforts were therefore initiated more than ten years ago to establish appropriate links with this cogeneric branch of industry.

The TAM tunnel support machine

Following a number of deployments involving compact machines of the type and size normally found in the coal industry the first truly “gigantic” GTA system to

attract worldwide attention was built and put into operation for support setting work in the St. Gotthard base tunnel in Switzerland. Even the professionals were amazed that 60 t of steelwork could simply be “hung from the tunnel roof” and that the colossal machine could then move about in any direction it wanted. Two even larger machines, each weighing about 80 t, are currently in service in a tunnelling project at Toulon in southern France, where they are not only setting the supports but also doing the drilling work too.

The TAM tunnel support machine is built up from the following components (Figure 2):

A mobile machine frame that is suspended from ñtwo supporting overhead monorail tracks.Four integral slider tracks to accommodate the ñ

individual appliances.Two drilling arms that can be deployed longitu- ñdinally for drilling the pipe screen and setting radial and roadhead bolts.

Figure 2. The TAM tunnel support machine.

http://www.gta-maschinensysteme.de/en/index.asp?p=116


MACHINERY

Two longitudinal working cradles that can move ñfreely and independently of each other.A support manipulator built into no. 1 cradle. ñTwo pipe racks flange mounted to the cradles ñand fitted with hydraulic pipe lifters.An optional shotcrete manipulator flange ñmounted to each working cradle.An electrohydraulic twin pump unit installed ñon a flange mounted support at the end of the machine frame. Two/four AKS 9300 drive trolley units. ñUser programmable radio remote control. ñ

The success of the two aforementioned machine de-ployments was enough to draw the attention of the international tunnelling community to the possibilities of using the “upper tier” in tunnel drivages – and this put the spotlight on the GTA system.

It has already been suggested that by bringing the upper working level into the equation through the introduction of the GTA system the conventional tunnel drivage can now be regarded as a further de-velopment of the New Austrian Tunnelling Method (NATM). With the NATM the crown section of the tunnel is excavated first and the lower section is then removed after a certain time lag – making it also a two-tier system. The new development lies in the fact that by employing GTA machine systems both upper and lower excavations can be undertaken as parallel operations. This reduces the cost of the project and makes for improved safety.

The Aachen-based engineering consultants WBI are currently conducting a comparative study of the time and cost expenditure incurred in completed tunnelling projects that have involved conventional tunnelling methods and “upper tier” systems.

Systems designed for the upper tunnel section have always been central to GTA thinking and will continue to be so in the years ahead. And GTA is fur-ther consolidating its market position by developing various other machines and systems for the tunnelling industry as a whole.

TEM tunnel enlargement machine

As well as producing machines for tunnel support work GTA has also been involved in developing technical solutions for the renovation and enlargement of old tunnel structures. In a pilot project undertaken for Deutsche Bahn, which was extensively reported in the press and media, two old small-section tunnels were renovated and extended without any interruption to train services. This was all made possible by GTA’s own tunnel enlargement machine, or TEM (Figure 3).

Figure 4. The GTA NormLifter tunnelling vehicle.

Figure 3. Design layout of the

TEM machine.


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This machine essentially comprises a protective por-tal (a tunnel within a tunnel) that is equipped with all the machinery and components needed to complete the widening of the original tunnel profile.

At about the same time the company developed and built a track removal machine for the renovation of the Arlberg tunnel in Austria. This mechanised system is able to replace both rail tracks in sequence while trains continue to operate on the free section of line.

Ancillary products

With a view to providing technical solutions for the demands of an extended market GTA has also in-vested real effort in developing a number of ancillary products. This included the take-over of the InTrak engineering firm (Innovative Traction Systems), a company with outstanding credentials as a manufac-turer of both overhead and floor-mounted transport systems (monorail installations and special climbing locomotives).

These innovative railborne traction systems for moving loads in tunnel construction projects – which can be diesel or electro-hydraulically powered – were developed and manufactured in Germany and have been used by the industry with great success both at home and abroad. InTrak also specialises in building machines to customer specifications that are designed for the automatic transport and transloading of mate-rials in inclines of up to 17°. The company’s portfolio also includes ATEX-approved diesel powered trolleys for the mining industry along with a range of hydraulic and pneumatic shunting trolleys.

GTA has therefore taken the first step to becoming a system supplier, particularly as far as the tunnelling industry is concerned.

Another key move in this direction was taken when GTA acquired 50 % of the German-French company NormRent, a machine rental business that has become a well established platform in the European market as a provider of mobile vehicles for the tunnelling industry. NormRent supplies equipment to tunnel contractors inside and outside Germany, including single- and twin-arm elevating platforms with on-board support and shotcrete manipulators and a range of mobile mixers and dumpers.

This means that GTA has both extended its inter-national client base and at the same time established a link to the type of tunnel-floor technology that has to operate in close conjunction with the company’s overhead systems.

In the course of this partnership it became apparent that the standard vehicles generally being bought in via NormRent did not meet the needs of the market in a number of respects. GTA therefore set about developing and building its own tunnelling vehicles in order to meet all the needs of this particular sector. A number of these machines have already gone into active service (Figure 4).

GTA is therefore well positioned for the future and will continue to provide its customers with a range of services based on the company hallmarks of innovative technology, efficiency and reliability.

http://www.olko.info


SHAFT SINKNG

Shaft sinking north of the Arctic CirclePlanning, construction and installation of the WS 10 mine project for MMC Norilsk Nickel, Moscow, RussiaDr.-Ing. Oleg Kaledin, Dipl.-Ing. Andreas Neff, Dipl.-Ing. Dietmar Schilling, Ing. Andre Marais, Thyssen Schachtbau GmbH, Mülheim an der Ruhr, Germany

Three years of tendering and negotiations came to an end in September 2007 when Thyssen Schachtbau GmbH was commissioned by OJSC

MMC Norilsk Nickel (MMC Norilsk Nickel) to under-take the “planning, construction and installation of the WS 10 ventilation shaft site for the Skalisti Mine”, which is part of the ore mining complex at Norilsk on the Taimyr Peninsula, a mining region of Siberia north of the Arctic Circle.

The contract effectively involves the planning and con struction of an ore mine and its integration and link-up with the existing ore production facility.

MMC Norilsk Nickel – company profile and history

MMC Norilsk Nickel is a Moscow-based russian mining and metallur gical company. The operational side of the business is centred in the northern Russian mining region of Norilsk-Talnach.

MMC Norilsk Nickel is the world’s leading producer of nickel and palladium. It also ranks among the largest international suppliers of platinum and is in the top-ten list of copper producers. In addition to this the company is involved in the mining of cobalt, rhodium, silver, gold, tellurium, selenium, iridium and ruthenium.

Though it was known as far back as the 17th century that there was copper and nickel on the Taimyr Penin-sula mining did not in fact commence in and around Norilsk until the 1920s. The USSR Government set up the Norilsk Combine on 23. June 1935, thereby laying the foundations for the world’s largest mining and

Thyssen Schachtbau GmbH accepted the challenge to sink a 2,055 m deep shaft with an inner diameter of 9.0 m for the worlds leading producer of nickel (MMC Norilsk Nickel) north of the arctic circle as a turn key project. Following a reduced workprogram by reasons of the finance crises in 2009, the operations are back in its entirety since January 2010. After completion of the 18 m deep shaft collar with fan drift and the core installations at the surface, the foreshaft sinking from 18 m level to a depth of 150 m can start in May 2010.

17th century Existence of copper and nickel deposits known

1920 Mining commences

1935 Norilsk Combine set up

1935 Severonickel Combine established in Monchegorsk

1940 Petchenganikel Combine set up in Zapolyarny

1989 The three Combines are merged with three other companies to create No-rilsk Nickel

1993 RAO Norilsk Nickel

2000 Norilsk Nickel Company

2001 Mining and Metallurgical Company: OJSC MMC Norilsk Nickel

Table 1. History of Norilsk Nickel.

metallurgical company for non-ferrous metals. The mine officially started production on 10. March 1939.

Combines for the production of copper and nickel were also established in the Murmansk mining region on the Kola Peninsula. The Severonikel Combine was founded at Monchegorsk in 1935 and the Petchen-ganikel Combine, which was based at the towns of Zapolyarny and Nikel, followed in 1940.

On 4. November 1989 a Government resolution merged the three combines with another two com-panies and the Gipronickel Planning Institute to form Norilsk Nickel. A further resolution of 30. June 1993 set up the joint-stock company RAO Norilsk Nickel. The company was then restructured in 2000 and re-named the Norilsk Mining Company. Further reorganisation took place the following year and as a result the group adopted its current name of “Mining and Metallurgical Company” (MMC Norilsk Nickel) (Table 1).

Shaft sinking north of the Arctic Circle

The Norilsk mining area is located in the Arctic Circle region of central Siberia to the east of the Yenisei River. At these latitudes the freezing temperatures that constantly prevail mean that the permafrost soil only thaws down to a maximum of 6 m for a short period during the two summer months. The Norilsk mining region contains rich ore bodies, with nickel, copper and platinum the main focus of mining activi-ties (Figure 1).

The WS 10 ventilation shaft is also to be constructed above an extremely rich body of ore and will form part of the new Skalisti mining development whose deposits will secure ore production from the Norilsk mining area from the end of 2015 well into the future. In addition, work will soon be commencing on the

http://www.thyssen-schachtbau.de/


SHAFT SINKING

SKS 1 production shaft, which is to be located about 1,500 m from the WS 10 site.

The two WS 10 and SKS 1 shafts for the future Skalisti Mine are about 35 km to the north-east of Norilsk and close to the urban district of Talnakh in the middle of the tundra. A 3 km-long paved road connects the WS 10 site to Skalisti’s WSS 7 main shaft, which has already been completed. Before the shaft construction work started Norilsk Nickel laid pipelines to supply the site with compressed air and fresh water and also installed an overhead power line for the supply of electricity.

The extreme climate conditions prevailing within the northern Arctic Circle posed a huge challenge for the construction crews (Figure 2). In winter temperatures in the Norilsk mining area are generally between –20 °C and –35 °C. However they can also plunge to below –40 °C and stay there for several days on end, making it impossible to operate load-lifting cranes, transport vehicles and other types of machinery. However the biggest problem is posed by the snowstorms that prevent any kind of outdoor work from being under-taken, and in some cases even make it impossible for heavy machinery to reach the worksite.

A logistic challenge

Supplying the construction site with machines and equipment is a huge logistic challenge. Deliveries from

Figure 1. Some-times a shaft will be established here, some 8 km away from the next living area deep in the tundra.

Figure 2. Clearing from snow in Norilsk official called „snow-fighting“.

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SHAFT SINKNG

Germany have to take the sea route from Rotterdam to Murmansk and then be transported onwards to Dudinka. As the ports of Murmansk and Dudinka are closed from the end of April to mid-July because of adverse weather conditions all supplies have to be carefully planned several months before they are shipped to the site and must also be registered well in advance with the customs authorities and shipping companies.

Materials supplied from within Russia, such as pipe-work, vehicles, items of equipment and concrete rein-forcement, can be transported on the Trans-Siberian Railway to Krasnoyarsk and from there shipped down the Yenisei River to Dudinka. These internal shipments also have to be planned in advance in great detail, as the Yenisei is only navigable during the ice-free months of July to September.

Scope of the contract

The volume of work assigned to Thyssen Schachtbau GmbH as the general contractor essentially involves the construction of a complete mining facility:

The surface work mainly comprised:

– Construction of an accommodation block for some 150 workers.

– Erection of the permanent 66 m-high shaft head-frame.

– Construction of the winder house and installation of two SIEMAG M-TEC2 twin-drum winding machines.

– Installation of two Howden mine ventilation fans delivering approx. 750 m³ of air per second, along with fan housings and diffusers.

– 110 kV and 6 kV AREVA transformer station for permanent power supply to the shaft.

– Workshops and stores buildings.

– Temporary concrete mixing plant and temporary fan building for the shaft sinking.

– Permanent belt installation to serve the waste tip.

Table 2. Surface structures.

Above ground the project includes the preparation ñof the shaft site, the erection of all temporary and permanent surface buildings and mine facilities and the installation of the shaft surface infrastructure (Figure 3, Table 2).Work to be carried out underground comprises the ñsinking of the 2,055 m-deep ventilation shaft, the construction of the fan drift and the excavation of shaft landings, pump rooms, roadway entries and a shaft undercut (Figure 4).

The preparatory work at the shaft site, which is located in a river valley in the foothills of a range of moun-tains, required extensive earth-moving operations to be carried out under strict environmental conditions in the natural landscape of the tundra, the aim be-ing to create an average gradient of 1.5° over the 40,000 m² site. The shaft surface was covered with a 1.5 m-thick layer of aggregate and surrounded by a drainage trench.

Because of the permafrost the foundations for the buildings had to stand on drilled piles 700 mm in diameter and set on compact rock. The piles were between 7 and 20 m in depth, depending on the local ground conditions at the shaft site.

The shaft, which is to be sunk by conventional drilling and blasting, has a specified finished diam-eter of 9.0 m and an end-depth of about 2,055 m. The 135 m-deep section of foreshaft has a concrete-backfilled cast-iron tubbing lining, while the main shaft section is to be constructed in steel concrete from the 135 m level to shaft bottom.

The blasting holes, which have a maximum depth of 4.5 m, are drilled with a pneu matic, six-arm shaft drilling machine. This rig can also be employed for large-profile holing work and for drilling exploration boreholes to a maximum length of 60 m.

The foreshaft is excavated using a portal crane and kibble system, with a hydraulic excavator loading out the debris on the sinking floor. The main shaft section will be constructed using a 7 m³-capacity dirt bucket and a grab with a payload of some 1.2 m³.

An innovative six-deck working platform system is used for installing the permanent shaft lining and shaft

Figure 4. Connec-tion of the water-proof fan drift is a first time used new concept for Norilsk.

Figure 3. The shaft facility in Septem-ber 2009; left to right: Basestruc-ture of the 110 kV power station, 45 m of final 66 m of permanent headframe, 6 kV station, temporary ventilation build-ing, temporary social building.


SHAFT SINKING

fittings, some of which are to be fitted in parallel with the sinking work. This six-deck platform moves on a walking mechanism without the need for the normal scaffold winches and is a very effective piece of equip-ment for the sinking of “ultradeep” shafts.

The two permanent SIEMAG M-TEC2 (today: SIEMAG TECBERG) twin-drum winding machines are available for the sinking operation. The kibble winch is essen-tially used for manwinding and debris clearance and for supplying materials to the sinking crew.

The SIEMAG M-TEC2 twin-drum winder, which will operate the skip conveyance when the mine comes into permanent production, is to be equipped with two manwinding cages during the sinking phase so as to keep the platform team supplied with materials such as pipework, shaft guides and buntons.

The concrete for the shaft lining is mixed above ground and then transported by a pump and drop-pipe system to the placement point behind the formwork at the level of the shaft platform, where it is compacted. A combination of rockbolts, wire mesh and, where required, shotcrete is used to provide temporary cavity support at the sinking floor. The shaft insets, pump rooms and roadway entries, along with the shaft bot-tom road, are all to be excavated conventionally by drilling and blasting with shotcrete spraying.

The entire shaft complex is to be handed over as a turnkey facility at the end of 2015. The TS pro-gramme of work has so far not deviated from the project schedule.

Progress to date

Earthmoving and ballast laying work commenced at the shaft site on 28. November 2007, which was just two months after the contract was awarded (Figure 5). In spite of the adverse weather conditions – the site was covered with up to 4 m of snow during the winter 2007/2008 – the earthmoving and excavation work progressed speedily. Some 150,000 m³ of tundra and loose soil were moved and about 60,000 m³ of crushed stone ballast laid in place. Preparation work at the shaft site was therefore practically completed by the autumn of 2008. A tipping point was also set up to accommodate the sinking debris.

After the local authorities had issued the regula-tory blasting permit the first round of blasting for the shaft profile commenced in July 2008. The shaft collar section was constructed to a depth of about 18 m and lined with cast-iron tubbing (Figure 6). The shaft cel-lar and four sets of headgear foundations were also completed during 2008 (Figure 7).

The excavation for the fan drift, which in some places was 16 m below ground level, was also constructed by drilling and firing. The drift broke through into the shaft in August 2008. The connection between the shaft and the fan drift was formed using an adapter frame, which was bolted up to the tubbing segments to create a watertight, flexible junction. The shaft cel-lar and entire length of the fan drift were lined with a waterproof welded membrane.

Work on the drilled piles for the accommodation block began in early May 2008. After the foundations had been completed and a lattice of steel beams

erected the actual building work was able to commence at the end of August. The accommodation block with its adjacent wastewater treatment plant will provide washroom facilities for about 150 workers (Figure 8). The unit also contains offices and conference rooms, along with lounge areas. The accommodation block was ready for occupation in February 2009.

Figure 7. Some meters ready sunk which was easier than to over come all that admini-strativ barriers.

Figure 6. Every shaft project

starts with a small step, here already four months after

contract awarding.

Figure 5. Surveying on difficult conditions with equipment up to date.


SHAFT SINKNG

The 6 kV power supply station temporarily required for the shaft sinking operation was fully installed and commissioned. The temporary fan building, a store-house and the permanent 110 kV and 6 kV transformer units are all currently under construction (Figure 9). The bottom 45 m section of headgear has been erected and is now being covered with steel sheet cladding. The complete headframe structure, which will stand 66 m in height, is to be ready by early 2010.

Outlook

Developments on the international financial markets, and the fall in the world market price for nickel, have forced MMC Norilsk Nickel to reschedule the WS 10 shaft complex project by gearing down the construc-tion work for the year 2009. The work is back in full operation since January 2010. The foreshaft sinking from the 18 m level to a depth of about 150 m can start in May 2010.

The installation of the permanent SIEMAG TECBERG twin-drum winding machines, which will be used to sink the WS 10 shaft from the 150 m level to its final depth of about 2,055 m, has now been postponed until August 2010.

Summary

In spite of the extremely exposed location of the shaft site good progress has been made on the extensive surface facilities, including the construction of the ac-commodation block, the shaft cellar, the 18 m-deep shaft collar, the fan drift and the permanent headframe structure (Figure 10).

The extreme climate has affected not only the planning of the project but also every aspect of the construction work and has required the planning engineers and site managers in particular to be able to adapt to events as they arise and to have the flex-ibility needed to acquaint themselves with building methods not known so far. Because of the limited transport options and restricted payload capacity the procurement and transport of equipment and materials, which have to be shipped via the Polar Sea or down the Yenisei River, has to be organised many months in advance and well before the items in ques-tion are required at the construction site.

The planning, construction and assembly work required for the WS 10 shaft complex has posed a number of exceptional problems. However, thanks to the commitment and technical skills of the construc-tion crews and the excellent working relationship established with the personnel from MMC Norilsk Nickel, all these difficulties have been successfully resolved and the various construction phases have been delivered on schedule.

Figure 10. Some 2,800 m³ concrete installation prepa-ration work for the shaft cellar in parallel to the final headframe installation.

Figure 9. Installa-tion of the shaft

building from which the highest

head-frame of Norilsk will be

built.

Figure 8. Good humor some

adverse conditions notwithstanding.


HEALTH AND SAFETY

A Safe Method for Cutting down Production Loss in Case of a Mine FireDr.-Ing. Walter Hermülheim, Central Department of Occupational Health, Safety and Environment, RAG Aktiengesellschaft, Herne, Germany

Dipl.-Ing. Andreas Betka, Central Mine Rescue Station, RAG Deutsche Steinkohle AG, Herne, Germany

During the first decades of fully mechanized longwall mining, as a matter of fact sponta-neous combustion fires could only be fought

under a high risk, e.g. by conventional sealing, by digging out or by measures of ventilation-pressure regulation. Today the classical cycle “Sealing – Wait-ing – Reopening” can be cut down to only a few days by means of an inertisation. Production can restart under safe conditions while fire fighting work is still going on.

The formerly high danger when sealing a fire can be reduced considerably provided the described safety regulations are observed. In this way a properly ap-plied inertisation contributes to cut down production loss in case of a mine fire, to preserve the coal deposit as well as to improve the safety of the mine rescue brigade itself.

Ignition Hazards

In the past spontaneous combustion fires in the goaf close to the longwall or the roadway have proven to be particular ignition hazards. Since the mid sixties there were over 200 spontaneous combustion fires with more than a dozen cases of CH4 ignitions.

Furthermore, in addition to smaller gas fires which normally can be safely put out by the workforce themselves, there were a few larger, open gas fires

Due to adequate preventive measures against fires and explosions, mine rescue work has become rare in the coal mining industry of the developed countries. The main field of activity of the rescue brigades today comprises the prevention of damage to property. Besides being responsible for the safety of the un-derground workforce, mine rescue brigades can contribute in this context reduce produc-tion losses. It should be agreeable that mine rescue work for the prevention of damage to property should not be more hazardous than regular underground work. This requires a regular check of safety regulations for riskre-lated work, e.g. as to high climatic loads or to the fighting of fire types which include the potential risk of an explosion. This paper points out the experiences of the German coal mining industry gathered on the latter subject during the last three decades. Special emphasize is put on ignition hazards related to spontaneous combustion fires.

which required action by the mine rescue and even inertisation in some cases (Figure 1).

Basic Rules Regarding Spontaneous Combustion Fires

The most fundamental safety rule when fighting spontaneous combustion fires in the goaf which are close to the longwall and to the roadway and thereby explosion prone (Figure 2) is to agree on a safe boundary value for retreat from the fire and for the last possible opportunity for a goaf or full space inertisation.

Based on the analysis of the ignition events men-tioned above, it has proven to be reasonable to set this boundary value at a fuel concentration of 50 – 60 % of the lower explosive limit in the air leakages escaping from the fire through the goaf. That means that organised work in roadways close to the seat of the fire in the goaf, e.g. sealing or the preparation

Figure 1. Mine Res-cue Training – Drill-ing into the Goaf.

Figure 2. Resealing a Roadside Pack with Sprayed Concrete.

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HEALTH AND SAFETY

mixture by a timely preparation of a goaf inertisation (Figure 4). Below of approximately 10 Vol.-% O2 the CH4 concentration is irrelevant (see Figure 3).

If in addition noteworthy concentrations of Hy-drogen and Carbon Monoxide appear in the combus-tion gasses, Coward-Diagrams for gas mixtures with several combustible components must be used for the evaluation mentioned above to reset the boun-dary values.

Coal Production during a Fire

If, besides the boundary values mentioned above, a number of other safety criteria are met, different from procedures in most other coal producing countries, production can safely be resumed even when the fire is still going on and is not yet fully extinguished. These rules include:

The location of the hidden fire is known or can ñbe estimated with sufficient precision.The goaf is separated from the open mine ñworkings and sealed by roadside packs which are constructed from solidly setting mortar matter.Known hollow spaces in the goaf behind the ñlongwall/ roadway crossing point are filled or interrupted by barriers constructed from mortar matter, e. g. when the ventilation “windows” in the roadside pack are filled after the face has advanced (Figure 5).The permitted boundary values for hazardous ñgasses are not exceeded and the Oxygen content remains above 18 % in the airflow within the open mine workings.Continuous monitoring of the air leakages ñwithin the goaf via ventilation pipes prove that there are no explosive gas mixtures present within the goaf close to the fire (negative ex-plosion index). A sufficient distance from the explosive limits must be kept at all times (see Figure 3).The Graham index (does not apply during iner- ñtisation) and the CO generation do not show a tendency to rise.

of an inertisation, is possible for as long as the CH4 concentration of the gases leaking from the fire meas-ured in the sniffing pipes which penetrate through the roadside pack into the goaf is lower than 2.5 to 3 Vol.-% CH4 (Figure 3).

Goaf Inertisation

Alternatively the Oxygen concentration in the air leak-ages can be kept at a safety margin of 2 % below the nose limit of the explosive triangle for the respective gas

Figure 5. Ventilation windows: Every 7 to 10 m ventilation windows are kept open in the road-side pack for climatisation and methane control purposes. After the face has advanced approximately 20 m past the window, it is sealed.

Figure 4. Goaf Inertisation.

Figure 3. Coward-Diagram for Methane with safety boundary to the explosion triangle (dashed) for fighting spontaneous combustion fires in the goaf.


HEALTH AND SAFETY

The fire does not generate more CO than the ñalarm value of 20 l/min. Deviations from this are possible, but they require the mutual consent of the head of the control centre and of the representatives of the central mine rescue sta-tion and the mines inspectorate within a risk assessment by the operation control.Goaf inertisation is applied if it is necessary to ñmeet the requirements described above.

It must be kept in mind that the rules mentioned above were in their full extend drawn up for goaf fires close to the longwall/roadway area, which are known to have caused ignitions in the past due to their proximity to sizable unventilated i. e. methane loaden cavities in the goaf.

Therefore a precise knowledge or an educated guess regarding the location of the fire is not always necessary, if nothing points to proximity of the event in the goaf to the longwall/roadway area and if the fire zone has definitely no connection to goaf cavities close to the longwall/roadway. If the distance from the face exceeds 300 m experience tells that no possible pathways for ignition do exist.

Full space Inertisation

If it becomes necessary to fully render the danger area inert (full space inertisation) from a safe distance due to a considerable risk of explosion, it has proven to be of advantage to improve the existing rule for isolating parts of the mine workings – construction of explosion proof seals around the affected area and subsequent simultaneous closing – by a stepwise approach and reduce the risk even further (Figure 6).

To do so, the airflow towards the fire is locked by means of a brattice at a safe distance from the danger zone and fed with Nitrogen. The return airflow from the fire is locked after the workings affected and the goaf have been flushed with Nitrogen for several times. A safe distance for inertisation of a possible ignition hazard is normally a kilometre between the place where a possible explosion is extinguished by the next explosion barrier and the brattice which cuts off the airflow to the fire. The brattices in the return airflow from the fire can be installed as close to the fire as possible, because they are constructed when the Oxygen concentration has dropped to a safe level.

Criteria which suggest that a full space inertisation is necessary are:

The presence of indicators pointing to a fire ñexisting already prior to a deflagration or ex-plosion (repeated ignitions are possible, maybe in intervals).CO concentrations which do not drop back to ñthe original value after a deflagration or explo-sion (repeated ignitions are possible),A gas fire which spreads in the goaf or other ñcavities and therefore or because of its general size cannot be controlled by direct approach,Dangers by a methane drainage pipe close to ñthe centre of the fire which cannot immediately be controlled.

To safe costs, it is advisable to have a goaf inertisation system operational before safety relevant boundary

values are exceeded. This way the expensive and potentially dangerous intermediate step “full space inertisation” can be omitted.

It must be stressed the full space inertisation does not require explosion-proof seals, rather light

Figure 6. Full Space Inertisation of a spontaneous combustion fire which poses an ignition hazard (top) and of an open fire (bottom).

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HEALTH AND SAFETY

Compared to the described inertisation procedure the classic sealing procedure without inertisation for the control of such incidents has almost completely disappeared. Figure 8 shows an explosion-proof seal constructed from mortar matter which will withstand static loads of 5 bar. Such seals are used to finally seal off abandoned mine workings in an explosion-proof manner.

Reopening the Fire Zone

The only way to make sure whether a fire has burned out or has been extinguished or contained via full space inertisation is to restore the ventilation in the area, which was rendered inert, on a trial basis.

This requires preparations to immediately reseal the opened brattices (air locks) and to immediately restart the full space inertisation if the fire surges up again. For such a trial ventilation the area around the affected zone should be cordoned off in ample distance. All activities are undertaken by the mine rescue under breathing and flame protection. If the ventilation trial must be aborted because the fire surges up again, the procedure and time scheme are the same as for a new full space inertisation.

Preparation and Precaution

On a case to case basis it is advisable to have a prepared goaf inertisation system which can be operated from a safe dis-tance in place before a possible incident to make a costly and elaborate full space inertisation unnecessary in the case of a explosion hazard. Such measures are especially advisable for coal seams which are prone to spontaneous combustion and have residual coal close to the roadway or the face in end position.

In the case of a positive Y ventilation system the use of so-called lost ducts in the salvaged area of the air-drawing roadway is an opportunity. Such ducts should be overlapping and coupled in such a way that that a tube ends respectively between 20 m and 100 m behind the face (Figure 9).

In seams which are prone to spontaneous combus-tion and in which shortly before the end position is reached, residual coal must be left in the goaf, it is advisable to install lost ducts between the shields on the fresh air side of the residual coal not closer than 10 to 15 m to the area of the salvage zone. When the end position is reached, the area behind the shields including the cavities on the edges of the longwall can be filled with foamed phenol resin.

References

1. Hermülheim, W. ; Beck, K.-D.: Inertization as Means for Reducing Down Time and the Explosion Risk in Cases of Spontaneous Combustion. Proceedings 6th Int. Mine Ventilation Congress. Pittsburgh, 1997.2. Hermülheim, W. et al.: Handbuch für das Grubenret-tungswesen im Steinkohlenbergbau (Colliery Mine Rescue Handbook, in German). Essen: VGE-Verlag, 2007.3. Hermülheim, W. ; Bolesta, M.: Examples of Grouting Techniques in Operational Applications. Proceedings 6th Int. Conf. on Rock Bolting & Injection Techniques and Roadway Support. RWTH Aachen Technical University, 2008.

Bild 9. Prepared Inertisation –

Overlapping lost Ducts in the

Salvage Zone of the Top Road.

Figure 8. Seal constructed from

Mortar Matter with all necessary

Fittings.

Figure 7. Water Explosion Barrier.

brattices are entirely sufficient. The explosion pro-tection is pro-vided by the Nitrogen and sufficient safety distances. The required safety distance of 1 km is from the spot at which a potential explosion is extinguished by a water barrier. A network of such barriers (Figure 7) therefore is a precondition for the approach described above.


HEALTH AND SAFETY

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Please send me copy/copies of Please send me copy/copies of Wall map, 100 cm x 86 cm Digital version on CD-ROMEuropean Gas Pipeline System European Gas Pipeline SystemISBN 978-3-86797-043-3 · € 129,- plus shipping rate ISBN 978-3-86797-059-4 · € 230,- plus shipping rate

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European Gas Pipeline SystemEdited by Hubertus SchöneichRevised and extended edition 2009.100 cm x 86 cm, in four colours. Scale 1: 8 333 000Price € 129,-ISBN 978-3-86797-043-3

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Montebruchstr. 2 · 45219 Essen · GermanyPhone: +49 (0) 20 54 / 9 24 -123E-Mail: [email protected]: www.vge.de

The new wall map, which has been updated using information generously provided by various companies, illustrates the key role that natural gas plays in supplying energy to Europe’s industrial nations. It shows the network of existing and planned gas pipelines in Europe together with the connections to the most important supplier regions in the CIS, the North Sea and North Africa.

To take account of the increasing economic importance of liquid gas the map now shows the location of LNG terminals along the coasts of Europe and North Africa. It also identifies all the current production areas along with their estimated gas deposits.

The relief background showing the main mountain ranges gives some idea of the difficulties involved in planning and constructing the different pipeline routes.

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PLANNING

Hard rock mining the opencast way versus a combination of opencast and deep mining methods An energy balance sheetDr.-Ing. Alexander Hennig, Chief Engineer, Professor Dr.-Ing. Christian Niemann-Delius, B.Sc. Thorsten Skrypzak, Institute for the Opencast Mining and Drilling (BBK III), RWTH Aachen University, Aachen, Germany

T he Federal Republic of Germany has access to a large number of shallow rock and stone deposits that contain large quantities of high quality

mineral. However, the ongoing extraction of these deposits means that the mining levels are becoming every deeper. This in turn makes for ever longer trans-port routes and results in a significant rise in energy consumption as the opencast lifecycle increases. This effect is especially apparent when using the shovel and truck method of winning, a non-continuous mining technique that is widely employed in the rock min-ing and quarrying sector. One way of reducing the high total energy demand is to modify the in-plant transport routes and transport procedures. It was against this background that an energy usage study was carried out of a combined surface and deep-mine transport system, comprising both roadway and shaft transport techniques, as part of a research project carried out by the Institute for the Opencast Min-ing and Drilling (BBK III) at RWTH Aachen, Aachen, Germany, the resulting process then being compared with the traditional opencast system of dumper truck haulage. The ultimate aim of the project was to find out whether, and to what extent, such a combined system could reduce energy consumption levels in the

The amount of energy consumed when cutting solid rock has a significant impact on the overall profitability of a projected mining operation, especially when energy prices are high. The raw-materials industry is therefore now more focussed than ever on optimising the winning process in energy usage terms. Against this background, both when planning new mines and developing existing facilities, factors such as the winning process and machinery combination are in-creasingly becoming the centre of attention; these systems may be completely new to the industry or may until recently only have been employed in other branches. For the hard rock cutting sector, which has always been based around opencast operations, this increasingly means coming to terms with the options and benefits of using methods borrowed from the deep mining industry.

rock and stone mining sector and therefore offer an alternative to conventional surface mining methods. Another aim of the investigation was to examine that point in time, in respect of the status of the opencast mine, at which it would be most logical in energy us-age terms to make the transition from one transport system to the other.

Strategy

The analysis was based on a project-specific model of a new mining undertaking. As the diverse and countless individual features often present in real mining areas, for example in terms of the geometry of the mineral beds, were deliberately omitted from the fictional deposits the opportunity presented itself to give a paradigmatic character to the project findings. Based on the depiction of the deposits and planned working faces in their respective feature sizes a performance-based selection was made of the various items of plant and machinery, whereby the equipment layout and performance was designed for the optimum effect. This was followed by a quantification of the energy demand from the individual working cycles. In order to examine the mining operation as a whole and to develop a meaningful reference value the energy consumption rates were based on each tonne of mineral extracted. This approach was then used to carry out separate analyses for the opencast mining operation using dumper trucks and for the combined mining system using underground mineral haulage techniques so that both systems could ultimately be matched one against the other.

Deposits and layout of the excavations

The body of deposits used for the analysis is 260 m in thickness and extends over an area measuring 300 m x 400 m. The rock strata are assumed to have an aver-age density of 2.8 t/m³. This gives a workable content of 31.2 Mill. m³ or 87.36 Mill. t.

The natural stone quarry is designed for an annual output of 1.5 Mill. t. The opencast site has a bench height of 20 m and a bench slope of 80°. The width of the intermediate berms is 3 m. The selected layout comprises 12 working levels with a total excavated volume of 18,575,000 m³ of saleable mineral.

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PLANNING

Mining by opencast methodsThe deposits are mined on a non-continuous basis using conventional drilling and blasting methods. This is currently the most commonly used method for excavating natural rock and stone, though the options for employing alternative winning techniques have been greatly extended in recent years as a result of new technical developments in this field. Drilling and blasting is suitable for mining rock and stone of any strength category. Slurry explosives are used so that there is no likelihood of problems arising due to the ingress of water into the shotholes. Some 41,000 m of shothole have to be drilled every year for the multi-row firing patterns with their 115 mm-diameter holes. A performance-related drilling machine can achieve an average drilling advance of 40 m an hour in the solid rock.

The blasted rock has an average size of 700 – 800 mm and this material is loaded on to the dumper trucks by a hydraulic excavator fitted with a bucket of 6.5 m³ capacity. The operating conditions for a loading machine of this size, which are governed for example by the properties of the material being loaded and the factors that apply during bucket discharge, can be categorised – in this particular case – as moderately heavy to heavy. The study also took account of the fact that free-steered loaders are frequently used for this

Sub-process/machine Calculated energy consumption in litres of diesel or in GWh electrical energy

Excavation by drilling and blasting

1.4 Mill. l

Loading by hydraulic excavator

3 Mill. l

Bulk transport by dumper truck

20.1 Mill. l

Primary crusher 7.6 GWh

Auxiliary equipment 0.8 Mill. l

Total 25.3 Mill. l +7.6 GWh

Table 1. Energy consumption rates for various sub-processes in the open-cast mining industry.

work. However for this particular mining scheme such an arrangement proves to be more energy consuming, as in view of the longer loading times the number of transport vehicles would have to be increased in order to achieve the same production level.

In order to quantify energy consumption during product haulage by dumper truck the average transport distance to the destination point was calculated for each working level and each run was then subdivided into uphill and downhill gradients and horizontal sections. The calculations were based on a gradient of 10 %. It was found that a medium-sized truck with a payload ca-pacity of 38 m³ was well suited to this particular mining concept in terms of both fuel consumption and loading time in conjunction with the hydraulic excavator. Ac-cording to the manufacturer’s data the deployment of a truck of this size and output range would result in an average consumption of some 148,000 l of diesel fuel a year. However, as the mining operation advances so the number of vehicles needed will increase. This fact is determined by the dynamics of the mineral extraction points and the situation will gradually change both as regards the horizontal transport distances and, more significantly, in terms of the difference in height that has to be overcome.

As well as the main opencast operations of excavat-ing, loading and transporting the mineral the analysis also included the energy consumption levels of the primary crusher. This was done for two reasons: firstly, the transition from rock winning to mineral process-ing, as far as the primary crushing stage is involved, is sometimes fairly fluid, and secondly, the primary crusher had to be installed as a central component of the transport chain when setting up the subsequent combination of surface extraction methods and un-derground transport systems. While the excavation machines would be operating on a single-shift basis, the crusher was to be set up for a two-shift routine. The crusher unit was designed to deliver a throughput of 375 t/h.

In addition to the different excavation, loading and transport machines directly involved in the winning operation an opencast mine will feature a number of auxiliary items of equipment that are needed to supply and back-up the sequence of operations. The type of equipment involved, and the scheduling of its deployment, will vary greatly depending on the condi-tions that exist at the working faces. In the example selected here the equipment employed for these particular services was assumed to have an energy consumption rate in the mid range.

The Table 1 shows the energy consumption rates for the opencast mining operation as calculated on the basis of the total operating time of the excavation project in question.

Using these figures the energy consumption level for the opencast mining operation can be calculated at approximately 0.51 l of diesel fuel per tonne of mineral produced. In actual practice consumption rates tend to lie between 0.5 and 1 l of fuel per tonne of mineral. The comparatively low amount of energy consumed during the opencast mining of stone from the workings under analysis can be attributed to the model-based character of the depicted deposits and

Figure 1. Percentage-based contribution of indi-vidual items of plant to overall energy consump-tion during opencast mining.


PLANNING

to the well coordinated combination of the selected plant and machinery (Figure 1).

Mining by a combination of methodsThe concept of using the following approach for mining the deposits is based on the idea that the continuous increase in energy consumption for truck haulage operations, which will result from the ever longer haulage distances, can be prevented by employing a transport route that will remain con-sistent throughout the entire life of the workings. Stone excavation will continue to be undertaken by opencast mining methods, i.e. drilling and blasting, while the mineral transport operation will use an underground roadway network running beneath the deposits (Figure 2).

The working cycle used for the actual mining opera-tion, and the energy consumed in the course of this process, will not differ in any way from the mining operations described above.

The mineral extracted from the floor of the opencast workings is loaded and carried by free-steered LHD (payload capacity 14 m³) to a glory hole located in the centre of the mine. This glory hole was constructed by raise boring. The surface area of the stoping levels decreases with increasing depth. The average distance to be covered by the LHD is therefore at its maximum on the first level, in this case 175 m. The stone is then tipped into the glory hole and delivered to the under-ground roadway.

While the surface mining work is organised on a one-shift basis, the underground transport operation and shaft haulage cycle will be a two-shift process. This will ensure that the required production levels are achieved and at the same time will help to reduce the dimensions of the underground installations and equipment. The deep workings can therefore also be made smaller in scale and the outlay on drivage and tunnelling work can be kept to a minimum. Conversely, however, this concept will require the excavation of underground bunker capacity. The first of the two in-termediate bunkers, each of which will be designed for

half a day’s output of 3,000 t, will receive the mineral as soon as it has been tipped into the glory hole. This is the only way in which the one-shift mining process can link up smoothly with the two-shift transport operation under way below ground.

The stone is drawn off from the bunker by a vibrating conveyor and delivered to the primary crusher, whose design and energy consumption rate is equivalent to that of the crusher unit installed in the initial mining operation.

The second bunker is located adjacent to the sur-face winding shaft. After passing through the primary crusher the mineral is transported along a conveyor road by a horizontal belt installation 210 m in length and with a belt width of 800 mm. It has been shown that belt conveyors are the most energy efficient transport system for operations of this type. As the conveyor road runs horizontally the output of the belt drive can also be kept relatively low, in this case a mere 18.5 kW. The second intermediate bunker is needed because it is not possible to transfer the min-eral directly from the continuous belt conveyor to the non-continuous one-skip shaft installation.

A measurement belt, which weighs out transport units of approximately 15 t of mineral for the shaft skip, has to be equipped for a drive output of 110 kW. However this belt does not operate continuously over the entire shift time, which means that energy con-sumption at this installation can be greatly reduced. The same applies to the second vibro-conveyor that draws the mineral from No. 2 bunker.

The underground workings are ventilated by a mine fan and two auxiliary fans installed in the road-ways. Because of the automated transport process and electrically powered transport systems the total ventilation requirement is merely 450 m³/min, which equates to a fan output of about 9 kW. The mineral winding shaft serves as the downcast shaft, while the return air exits via a ventilation borehole that is combined with the surface glory hole.

The winding shaft is 345 m in length and has a diameter of 6 m. One shaft compartment is used for

Figure 2. Schematic representation of the underground roadway layout.


PLANNING

tion values established for the various part-processes and items of plant over the entire lifespan of the project.

The application of this particular winning method therefore results in an energy consumption rate of 0.67 l of diesel fuel (converted figures) per tonne of mineral extracted (Figure 3).

Analysis and assessment of the results

An individual assessment of the two working methods results in an energy consumption figure of 0.1 l/t for the conventional opencast mining technique and 0.67 l/t for the combined system, as based on the production of 1 t of mineral. The opencast method therefore works out as more energy efficient when viewed over the entire life cycle of the operation. The main reason for this result is the high annual energy demand of the shaft winding installation, which re-mains constant over the entire mining period, as the annual output of the mine does not change and is extracted from a consistent depth.

In the case of the dumper truck transport system as used for the opencast mining operations, on the other hand, there is a significant change in energy expenditure when taken over the entire life cycle of the project. As the winning depth increases there is a substantial rise in energy consumption and in the number of vehicles deployed. On the first working level, however, haulage from the extraction point to the reference destination only means negotiating a height difference of some 20 m. When the material is being transported along the underground roadways, on the other hand, and even when mining is still under way on the first level, it still has to be raised up the shaft from a depth of 345 m, after having been pre-viously tipped down the glory hole. This means that the shaft winder has a much higher energy demand than the truck haulage system.

Further investigations were carried out in order to determine at what working level, and hence at what mining depth, the opencast method being used in this particular mining scenario becomes more energy intensive than the underground conveyor system. Although the shaft winding installation has a high energy requirement, this is offset by the relatively low energy demand calculated for the rest of the combined underground transport system based on belt convey-ors. Even the roadway development work proved to be less energy intensive than had been expected. The Figure 4 shows the annual energy consumption for a mineral yield of 1.5 Mill. t from the individual working levels for both the opencast mining system and the combined system (Table 3).

It becomes clear that the combined system only becomes more energy efficient than the opencast method when mine level 10 is reached. Prior to this both energy consumption rates gradually move into line with each other. While the truck haulage operation becomes increasingly energy intensive, the amount of fuel consumed by the combined system’s free-steered ”load and carry” vehicle drops slightly as the mining levels become narrower. This is also reflected in the

Sub-process/machine Calculated energy consumption in litres of diesel fuel or in GWh of electrical energy

Development work 0.065 Mill. l

Excavation by drilling and blasting

1.4 Mill. l

Free-steered loader in load-and-carry mode

7.6 Mill. l

Underground transport equipment (vibro-conveyor, belt conveyor, measurement conveyor)

23.2 GWh

Primary crusher 7.6 GWh

Shaft winder 122.1 GWh

Ventilation (main fan and auxiliary fans)

2.5 GWh

Total 9.065 Mill. l + 155.4 GWh

Table 2. Energy consumption rates during individual sub-processes based on equipment combinations.

winding men and materials, while the second houses the skip conveyance. Because of the excellent stability of the local strata the shaft wall can be supported by a combination of rockbolts and wire mesh. This means that practically the entire shaft diameter of 6 m will be available as usable space. In order to achieve a winding speed of 12 m/s the shaft winder has to be designed for a skip payload of nearly 16 t and a drive output of 1,100 kW.

Before regular production can begin the calculations also have to take account of the energy consumed by the underground tunnelling work (conventional drivage system) and the dumping of the excavated debris at a spoil tip some 300 m distant from the shaft mouth. Both diesel-powered and electrically driven machines were used for the development work. For reasons of clarity the energy consumption levels of the different items of equipment were consolidated into a single value. This means that the equivalent diesel consumption rates were used for the electric-driven machines. The Table 2 shows the energy consump-

Figure 3. Percentage-based contribution of indi-vidual items of plant to overall energy consump-tion during underground operations.


PLANNING

overall analysis of the energy expenditure report. The further the mining operation progresses beyond level 10, which in this particular scenario represents a depth of some 200 m, the greater is the energy-consumption advantage of the underground transport system (Figure 5).

In simple terms this result means that the final depth of 260 m selected for this particular mining project was not large enough to fully exploit the benefits of the underground transport method. The additional energy expenditure incurred in the early stages of the underground operation, compared with that used by the opencast system, could not be compensated for as the project progressed from mine level 10 to its final depth. Even though the opencast transport lines became longer than the underground conveyor roads once the project reached level 3 (see Figure 5), it would still be some time before the shaft winding system could reach its energy ”break even” point.

The shaft winding machine is without any doubt the largest energy consumer in an underground mine. Mine fans also consume similar quantities of energy, but only at the larger mines and when air-flow require-ments are high.

Conclusions

The combination system being studied in the case of the selected area of deposits is overall more energy intensive that the conventional opencast extraction method. However, it is clear that once an opencast mine reaches a certain working depth, which would have to be determined on an individual basis, there are energy benefits to be gained from transporting the mineral along an underground roadway connect-ing to a surface shaft.

The certainty of this statement can primarily be attributed to the fact that neither the underground

Mine level

Energy requirement in litres of diesel fuel for a 1.5 Mill. t yield of mineral

Opencast Process combination

1 450,000 1,035,000

2 515,000 1,030,000

3 585,000 1,025,000

4 650,000 1,020,000

5 730,000 1,015,000

6 790,000 1,010,000

7 860,000 1,005,000

8 920,000 1,000,000

9 990,000 995,000

10 1,070,000 990,000

11 1,120,000 985,000

12 1,240,000 980,000

13 1,300,000 970,000

Table 3: Energy consumption for a mineral yield of 1.5 Mill. t.

development work nor the ventilation system has a significant influence on overall energy consumption and can quickly be compensated for.

Moreover, the network of underground roadways can still be constructed even after the opencast mining operations have commenced.

As the underground development work is being undertaken beneath the deposits there is no risk of the tunnel drivages colliding with the advancing stopes and benches above ground. This means that the limits of workability of a valuable mineral can be extended quite significantly to include deeper working levels, which makes for a considerable increase in the volume of recoverable deposits. For one thing, this can have enormous benefits when carrying out an economic feasibility study of an existing mining project, as de-posits will now become accessible that were previously not included at the planning stage.

For another, deposits that were previously consid-ered to be unprofitable can become exploitable. This makes a real contribution towards meeting future demand for raw materials.

Figure 4. Graphic representation of the energy requirement for the annual mineral yield from the individual working levels for both the opencast and the underground systems.

Figure 5. Distances to be overcome by the mechanised haulage/convey-ing systems operating between the extraction site and the reference unloading point


METHANE

International CMM and VAM Project Development Experiences of a Mining ConsultantDipl.-Ing. Thomas Imgrund, DMT GmbH & Co. KG, Essen, Germany

Most coal deposits contain more or less methane, which is adsorbed in the coal matrix and also occurs within pores, cleats

and voids of both, coal and surrounding rock. This methane is released as a result of mining coal and subsequent pressure release in the surrounding strata. If possible, methane emissions will be con-trolled by diluting and discharging the methane through the mine ventilation. If this is not suffici-ent, gas drainage will be carried out. Apart from being a safety risk in the mine operation, methane from coal mines is an energy source which could be utilised by various technologies for heat or po-wer generation (Figure 1). Regarding the climate, methane is also identified as a greenhouse gas, which is 21 times more harmful to global warming compared to carbon dioxide (1).

As in most cases the utilisation of methane from coal mines is not economic at actual prices for gas, electricity or heat, projects are often financed by the generation of carbon credits. This business is based on the Clean Development Mechanism (CDM) and Joint Implementation (JI) of Kyoto Protocol. These so called project based mechanisms allow emitters of carbon dioxide fulfilling their emission reduction

Methane is a waste product of mining coal. Utilisation of this gas is a more and more interesting issue for coal mine operator as well as a young business based on the carbon market. Apart from using drained gas for power generation using internal combustion engines or supply to the natural gas grid, thermal oxidation of highly diluted methane in the ventilation air is in the focus of interest. Projects for utilisation of this waste product are usually financed by carbon credits. The German mining engineering and consulting group DMT supports international companies during development of gas utilisation projects around the world.

targets by carbon offset projects in foreign countries. Carbon credits could be generated either by methane utilisation or just conversion of methane to water and carbon dioxide as a less harmful greenhouse gas. The abatement of 1 t of methane is equal with avoiding 18.25 t of carbon dioxide.

The methane discharged by ventilation (Figure 2) is called ventilation air methane (VAM) and the drained methane is called coal mine methane (CMM). Usually the concentration of VAM in the ventilation air is limited to 1 % maximum, depending on the individual national mine safety regulations. Other limits are common for particular parts of the mine like return airways of panels. The utilisation of such highly diluted methane-air mixtures has become technically and economically interesting by new technologies as thermal oxidation during the last years. Concentrations of CMM are considerably higher, but limits differ from country to country.

After the closure of a coal mine, methane emissions abate more or less quickly, but use to continue on a lower level for decades. This source is called abandoned coal mine methane (AMM). Methane concentrations of AMM depend on the residual gas contents after mining as well on the inflow of air into the disused mine workings. Methane could also be produced from virgin coal beds, as a source of natural gas or during pre-dainage of coal seams to be mined. This is known as coal bed methane (CBM).

Development of methane utilisation projects

Utilisation of coal mine methane has a very long tradition in Germany, but development was acceler-

Figure 1. CMM fired power plant

at a Chinese colliery.

Figure 2. Ventilation air as

an energy source.

http://www.dmt.de/index.php?id=11&L=1


METHANE

ated by the Renewable Energy Law in the year 2000. This has been supporting CMM and AMM to power projects with access to the grid and guaranteed prices for power input. Today there is more than 220 MW electrical capacity, installed at 128 CHP plants on dif-ferent locations. During the last 10 years, DMT GmbH & Co. KG, Essen, Germany, has played a major role in building up Germanys CMM and AMM industry and also has been involved in AMM and CBM projects in other European countries.

Based on this development, during the last years, DMT has cooperated with different parties involved in international carbon financed CMM and VAM projects. Clients are power generation companies, energy trad-ers, investment companies and developer of emission reduction projects. So far, DMT has been involved in about 65 projects worldwide, mainly in PR China, followed by Russia and Eastern European countries. In PR China projects in Heilongjiang, Inner Mongolia, Ningxia, Shanxi, Anhui, Chongquing, Yunnan and Guizhou provinces have been assessed (Figure 3).

It must be considered, that the stakeholders in CMM and VAM projects have different views on methane emissions. For the mine operator coal mining is the core business. CMM and VAM are waste products at first, produced by measures for effective and safe coal production. Mine ventilation and gas drainage have to meet the mine operation’s requirements and national mine safety regulations.

The project developer, as well as an investor or a buyer of generated carbon credits, is interested in the converted methane volumes. After investment in utilisation or abatement plants, the revenues from car-bon credits are essential for them. Maximum utilisable methane volume flows and adequate concentrations are the important issues.

As both parties have different interests or at least rankings of their interests, the common interests have to be elaborated as a base of cooperation. Anyway, the gas utilisation has to be managed after achieving coal production and mine safety. DMT understands its position as a bridge between these two parties, bringing in know how regarding the interactions between coal mining and gas emissions on one side and the requirements as well as the possibilities of gas utilisation on the other side.

Barriers in CMM and VAM projects

Due to the greenhouse warming potential of methane, CMM and VAM projects have a large potential com-pared to other emission reduction projects. However, since 2007 only 26 projects have been registered as CDM projects, all of them in PR China. A small amount was registered as JI projects in Eastern European countries. Among others, DMT indicates some major barriers in project development.

As mentioned, for the mine operator coal produc-tion is the main business. Hence, gas utilisation is of secondary interest. The following example shows the value of emitted methane compared to the value of mined coal. The example is a gassy mine in Eastern Europe producing 4 Mill. t of coal and emitting 118 Mill. m³ of methane annually. At a coal price of

130 US-$/t and a carbon credit price of 15 US $/t car-bon dioxide equivalent, the value of the total emitted methane is only 4 % of the value of the mined coal. As an abatement rate of 100 % usually is not economic, the possible revenues from carbon credits are even less, about 15 Mill. US-$/a. Nevertheless, this is recognized as additional business by many mine operators.

Further barriers are fluctuating methane volume flows and concentrations of both, CMM and VAM. This results in an utilisation ratio of considerably less than 100 %. At first, volume flows and concentra-tions depend on the mining and drainage methods. They could fluctuate short term, within hours, days or weeks, and long term, within months or years. Long term fluctuations could be a result of a changing coal production level as well as a result of increasing or decreasing gas contents within the deposit. Short term fluctuations are influenced by the current coal production, the operation of the drainage system and the barometric pressure.

Gas emission forecasts are carried out by mining companies or national institutes for dimensioning ventilation and gas drainage. Ventilation and gas drainage is designed based on the maximum expected methane emissions which have to be controlled dur-ing mining. In some cases they are compiled just for approval procedures without any interest in accuracy. They could differ from reality considerably. At a major mine DMT’s technical team has seen plans for a CMM to power plant with a gas demand exceeding the total gas emission of the mine, drained and vented methane. Misinvestment would have been more than 30 Mill US-$. Hence, for methane utilisation, methane emissions must be assessed in another way (Figure 4).

Consultancy services

DMT offers various services during project develop-ment and focus on bringing in special knowledge and

Figure 3. DMT’s activities in CMM and VAM project development in PR China.


METHANE

engineers as well as inspections of relevant surface and underground facilities (Figure 7).

For the analysis of statistic methane emission figures it is essential to understand how these figures were produced. Hence, it is necessary to discuss the measure-ment and documentation methods with the people who compiled these figures. For the mine operator, these statistics are often of limited interest. Electronic state of the art sensors are not unusual even at less developed mines, but sometimes no calibration is car-ried out due to additional costs. Hence, at some mines the figures produced by old fashioned measurement devices are more accurate than the figures measured by state of the art sensors. Nevertheless official methane emission statistics may base on electric sensors.

Gas content figures are a basic figure for gas emission forecasts. Hence, usually underground coal samples are taken for gas content determination in DMT’s laboratory.

The gathered data are analysed as a base for an independent forecast of total and utilisable methane emissions. As methane emissions may differ depend-ing on the mine district, the mined seam as well the mining, ventilation and drainage method, it is neces-sary to analyse gas emission figures differentiated. Therefore, the historic methane emission figures are broken down as detailed as reasonable. By this, the particular sources of gas emissions – the individual productive or abandoned mine districts, headings and panels and within this source the working seam and the particular surrounding strata – are identi-fied. The influences of coal production, ventilation and gas drainage operations on methane quantities and qualities are evaluated based on this detailed gas emission balance.

Gas emission forecast

Especially for VAM projects, the whole mine operation and the future mine development have to be analysed carefully, as small differences in VAM concentration could result in considerably misinvestment.

DMT compiles gas emission forecasts based on an empirical model developed during the last decades for the prediction of roof and floor gas emissions of high productive longwall panels. As the origin of the method is in German collieries under special conditions, it has to be adjusted and calibrated when used in other deposits. Scientific numerical simulations are only as accurate as the input parameter. These parameters like permeability of virgin and destroyed coal and rock are commonly not available. Hence, DMT uses a pragmatic forecast method based on the theoretical calculation and the detailed analysis of recorded actual and historic methane emissions.

The Figure 8 shows the forecast for a CMM project at a Chinese mine, working three seams successively from the top to the bottom seam. After mining the top seam the gas contents in the roof and the floor of the middle seam are reduced, so gas emissions will be considerably lower.

Deviations of actual from forecasted gas emissions are common and recognized worldwide. This applies also to the particular national forecast methods. As a

experience wherever it is reasonable. In the early stage of project development, DMT assists with directed data acquisition and interpretation with the target of separating economic from obviously uneconomic projects as early as possible. It must be considered, that mining operations, especially in PR China, cover a wide range of technical standards (Figures 5 and 6) and geological conditions.

The second step is a technical due diligence, includ-ing data acquisition and discussions with the mine’s

Figure 4. Dimensioning of CMM drainage and CMM utilisation.

Figure 5. Colliery at Western China.

Figure 6. Small coal mine in Northern China.


METHANE

result it is necessary calibrating the forecast with the reality very carefully for each project.

The methane emission forecast and the possibilities for energy supply lead to the methane utilisation con-cept. In most cases gas engines and thermal oxidation plants are the core of the concept. Apart from this, there are further possibilities for methane utilisation and abatement, for example direct gas distribution via pipeline injection, conventional boilers and CMM flares.

Risk assessment

The CMM or VAM project is reliant to the mine as gas producer. The project could depend on the advance of just one panel. This results in significant risks as the project developer, investor or credit buyer have almost no influence on gas delivery.

An omnipresent risk is a production reduction or even stop. Without coal production methane emis-sions in most cases decrease to near zero. Production reductions or downtimes could be a result of geologi-cal problems or gas outburst. DMT’s team has visited a mine construction site in Asia, which was stopped for about one year due to a major gas outburst. After this period mine development moved to another, less explored part of the deposit - with leaving the future quite uncertain.

Last but not least, gas quantities and qualities depend on the actual operation of ventilation and drainage and, of course, of the individual people and their understanding of the job. This includes the whole involved staff from the management to the working force actually drilling and operating the drainage boreholes.

Improvement of gas drainage and ventilation

The evaluation of improvement potential has three main targets: reducing the identified project risks

where possible, enhancing utilisable gas volume flows and enhancing coal production and mine safety.

The first possibility is adjusting the utilisation concept to the minimum gas delivery. That is more reasonable than building up a power plant and search-ing for additional gas afterwards. Flexibility might be important as no mine can guarantee the gas delivery for a whole project lifetime.

At CMM projects maximum and stable methane volume flows are of interest. Apart from an optimised borehole layout and adequate borehole sealing, the management of the boreholes is a solution for this.

At VAM projects an enhancement of concentrations is not trivial, as safety in all parts of the mine must be guaranteed. Overall reduction of ventilation air flow is no solution and highly dangerous. Nevertheless at many mines for example needless air shortcuts could be reduced. Apart from improvement of ventilation, CMM gas can be added to the oxidation plant.

Low concentrations of CMM are common in some countries, even in the explosive range. Apart from the safety risk, air shortcuts result in a reduction of drain-age system capacity. Reducing shortcuts is interesting for both, the CMM power plant operator and the mine operator, as this affects the operation costs of the drainage system.

The ventilation layout of panels affects the drainage ratio due to the possible drilling locations and the influ-

Figure7. CMM pumping station at a colliery in Eastern Europe.

Figure 8. Forecast of drainable and utilisable CMM.


METHANE

ence of the pressure drop of the ventilation. At retreat longwall panels drainage boreholes are often drilled from the gate road in advance of the coal face. These boreholes are frequently destroyed by the advancing coal face. Horizontal boreholes drilled from raises (as carried out in PR China) are more effective, but also more expensive. In the UK cross measure boreholes are drilled in the rear of the coalface of retreat long-wall operations. Ventilation of the drilling location is carried out by a back return method.

Advanced longwall operations with Y-form ventila-tion could be provided with more ventilation air. At panels with very high gas emissions or working thin seams this is essential, as the coal face width is a limit for the ventilation air flow. Due to the pressure drop of ventilation, maximum gas emissions are not con-centrated at the return air side of the coal face, but are dispersed within the return airway. Gas drainage is supported by the pressure drop and a full access to all boreholes. By regulation of each borehole the operation lifetime of the boreholes is increased and methane concentrations could be enhanced.

Specifics of VAM projects

Compared to CMM utilisation, VAM projects are considerably more ambitious. The development of future VAM concentration is essential, because of the technical limit of the oxidation plant (about 0.2 %) and a sufficient methane throughput. At a constant utilised ventilation air flow, the methane throughput depends on the concentration only.

At high mixture volume flows, a deviation in methane concentration measurement of only 0.1 % has an enormous influence in the total VAM volume. On the other side this is a tolerated deviation, which can be found at every mine. Therefore, the assess-ment of historic and actual gas emission data, as well as the methane forecast, have to be very accurate. Future variations of ventilation air flow have to be considered.

The example in Figure 9 shows the forecast of VAM concentrations for two cases. Currently at this mine there are considerably air shortcuts within the upcast

shaft, which are one third of the mine districts total airflow. These could be reduced with low cost im-provement by about 75 %. Without improvement the project would not be economic. Apart from additional revenues from carbon credits, the energy demand of the main fan could be reduced considerably.

Summary

Methane emissions at coal mines are not only a safety issue, but could mean additional revenues by power and heat production or the generation of carbon credits. With the Kyoto Protocol the utilisation or just abatement of methane emissions from coal mines has become a worldwide growing business. Nevertheless, for a mine operator methane is a waste product as first, while for the CMM or VAM project developer methane is the essential source for project revenues. During project development, these different points of view have to be considered. If not, the project won’t be sufficient for both parties.

DMT has been involved in the development of several projects for the utilisation of drained coal mine methane (CMM) and ventilation air methane (VAM), most of them in Asia and Eastern Europe. Within a technical due diligence, an independent forecast of utilisable methane emissions and an assessment of risks and improvement potential is elaborated. This is the base for investment in methane utilisation equipment. Misinvestment, caused by incorrect interpretation of gas emission statistics, gas emission forecasts and future mine development, could be avoided.

Improvement of both, gas drainage and mine venti-lation could be a benefit for both, the mine operation and the methane utilisation. Within the project, coal production and mine safety must be considered as most important. Hence, the different requirements of mining and methane utilisation have to be evaluated and common interests have to be elaborated for a successful project.

References1. Intergovernmental Panel on Climate Change (IPPC): IPCC Fourth Assessment Report (AR4), 2007.

Figure 9. Forecast of VAM concentrations.

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